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diff --git a/old/63784-0.txt b/old/63784-0.txt deleted file mode 100644 index b0ff346..0000000 --- a/old/63784-0.txt +++ /dev/null @@ -1,11942 +0,0 @@ -The Project Gutenberg EBook of Lead Smelting and Refining, by Various - -This eBook is for the use of anyone anywhere in the United States and most -other parts of the world at no cost and with almost no restrictions -whatsoever. You may copy it, give it away or re-use it under the terms of -the Project Gutenberg License included with this eBook or online at -www.gutenberg.org. If you are not located in the United States, you'll have -to check the laws of the country where you are located before using this ebook. - -Title: Lead Smelting and Refining - With notes on lead mining - -Author: Various - -Editor: Walter Renton Ingalls - -Release Date: November 16, 2020 [EBook #63784] - -Language: English - -Character set encoding: UTF-8 - -*** START OF THIS PROJECT GUTENBERG EBOOK LEAD SMELTING AND REFINING *** - - - - -Produced by deaurider, Les Galloway and the Online -Distributed Proofreading Team at https://www.pgdp.net (This -file was produced from images generously made available -by The Internet Archive) - - - - - - Transcriber’s Notes - -Obvious typographical errors have been silently corrected. Variations -in hyphenation other spelling and punctuation remains unchanged. In -particular the words height and hight are used about equally. As hight -is a legitimate spelling, it has not been changed. - -Some of the larger tables have been re-organised to improve clarity and -avoid excessive width. - -The footnotes are located at the end of the book. - -Italics are represented thus _italic_. - - - - - LEAD SMELTING - - AND - - REFINING - - WITH SOME NOTES ON LEAD MINING - - - EDITED BY - WALTER RENTON INGALLS - - - [Illustration: Publisher’s Device] - - - NEW YORK AND LONDON - THE ENGINEERING AND MINING JOURNAL - 1906 - - - COPYRIGHT, 1906, - BY THE ENGINEERING AND MINING JOURNAL. - - ALSO ENTERED AT - STATIONERS’ HALL, LONDON, ENGLAND. - - ALL RIGHTS RESERVED. - - - - - PREFACE - - -This book is a reprint of various articles pertaining especially to the -smelting and refining of lead, together with a few articles relating -to the mining of lead ore, which have appeared in the _Engineering and -Mining Journal_, chiefly during the last three years; in a few cases -articles from earlier issues have been inserted, in view of their -special importance in rounding out certain of the subjects treated. -For the same reason, several articles from the _Transactions_ of -the American Institute of Mining Engineers have been incorporated, -permission to republish them in this way having been courteously -granted by the Secretary of the Institute. Certain of the other -articles comprised in this book are abstracts of papers originally -presented before engineering societies, or published in other technical -periodicals, subsequently republished in the _Engineering and Mining -Journal_, as to which proper acknowledgment has been made in all cases. - -The articles comprised in this book relate to a variety of subjects, -which are of importance in the practical metallurgy of lead, and -especially in connection with the desulphurization of galena, which is -now accomplished by a new class of processes known as “Lime Roasting” -processes. The successful introduction of these processes into the -metallurgy of lead has been one of the most important features in -the history of the latter during the last twenty-five years. Their -development is so recent that they are not elsewhere treated in -technical literature, outside of the pages of the periodicals and the -transactions of engineering societies. The theory and practice of these -processes are not yet by any means well understood, and a year or two -hence we shall doubtless possess much more knowledge concerning them -than we have now. Prompt information respecting such new developments -is, however, more desirable than delay with a view to saying the -last word on the subject, which never can be said by any of us, even -if we should wait to the end of the lifetime. For this reason it -has appeared useful to collect and republish in convenient form the -articles of this character which have appeared during the last few -years. - - W. R. INGALLS. - - AUGUST 1, 1906. - - - - - CONTENTS - - - PART I - - NOTES ON LEAD MINING - PAGE - - SOURCES OF LEAD PRODUCTION IN THE UNITED STATES (WALTER - RENTON INGALLS) 3 - - NOTES ON THE SOURCE OF THE SOUTHEAST MISSOURI LEAD (H. A. - WHEELER) 10 - - MINING IN SOUTHEASTERN MISSOURI (WALTER RENTON INGALLS) 16 - - LEAD MINING IN SOUTHEASTERN MISSOURI (R. D. O. JOHNSON) 18 - - THE LEAD ORES OF SOUTHWESTERN MISSOURI (C. V. PETRAEUS AND - W. GEO. WARING) 24 - - - PART II - - ROAST-REACTION SMELTING - - SCOTCH HEARTHS AND REVERBERATORY FURNACES - - LEAD SMELTING IN THE SCOTCH HEARTH (KENNETH W. M. MIDDLETON) 31 - - THE FEDERAL SMELTING WORKS, NEAR ALTON, ILL. (O. PUFAHL) 38 - - LEAD SMELTING AT TARNOWITZ (EDITORIAL) 41 - - LEAD SMELTING IN REVERBERATORY FURNACES AT DESLOGE, MO. - (WALTER RENTON INGALLS) 42 - - - PART III - - SINTERING AND BRIQUETTING - - THE DESULPHURIZATION OF SLIMES BY HEAP ROASTING AT BROKEN - HILL (E. J. HORWOOD) 51 - - THE PREPARATION OF FINE MATERIAL FOR SMELTING (T. J. GREENWAY) 59 - - THE BRIQUETTING OF MINERALS (ROBERT SCHORR) 63 - - A BRICKING PLANT FOR FLUE DUST AND FINE ORES (JAS. C. BENNETT) 66 - - - PART IV - - SMELTING IN THE BLAST FURNACE - - MODERN SILVER-LEAD SMELTING (ARTHUR S. DWIGHT) 73 - - MECHANICAL FEEDING OF SILVER-LEAD BLAST FURNACES (ARTHUR S. - DWIGHT) 81 - - COST OF SMELTING AND REFINING (MALVERN W. ILES) 96 - - SMELTING ZINC RETORT RESIDUES (E. M. JOHNSON) 104 - - ZINC OXIDE IN SLAGS (W. MAYNARD HUTCHINGS) 108 - - - PART V - - LIME-ROASTING OF GALENA - - THE HUNTINGTON-HEBERLEIN PROCESS 113 - - LIME-ROASTING OF GALENA (EDITORIAL) 114 - - THE NEW METHODS OF DESULPHURIZING GALENA (W. BORCHERS) 116 - - LIME-ROASTING OF GALENA (W. MAYNARD HUTCHINGS) 126 - - THEORETICAL ASPECTS OF LEAD-ORE ROASTING (C. GUILLEMAIN) 133 - - METALLURGICAL BEHAVIOR OF LEAD SULPHIDE AND CALCIUM SULPHATE - (F. O. DOELTZ) 139 - - THE HUNTINGTON-HEBERLEIN PROCESS (DONALD CLARK) 144 - - THE HUNTINGTON-HEBERLEIN PROCESS AT FRIEDRICHSHÜTTE (A. - BIERNBAUM) 148 - - THE HUNTINGTON-HEBERLEIN PROCESS FROM THE HYGIENIC STANDPOINT - (A. BIERNBAUM) 160 - - THE HUNTINGTON-HEBERLEIN PROCESS (THOMAS HUNTINGTON AND - FERDINAND HEBERLEIN) 167 - - MAKING SULPHURIC ACID AT BROKEN HILL (EDITORIAL) 174 - - THE CARMICHAEL-BRADFORD PROCESS (DONALD CLARK) 175 - - THE CARMICHAEL-BRADFORD PROCESS (WALTER RENTON INGALLS) 177 - - THE SAVELSBERG PROCESS (WALTER RENTON INGALLS) 186 - - LIME-ROASTING OF GALENA (WALTER RENTON INGALLS) 193 - - - PART VI - - OTHER METHODS OF SMELTING - - THE BORMETTES METHOD OF LEAD AND COPPER SMELTING (ALFREDO - LOTTI) 215 - - THE GERMOT PROCESS (WALTER RENTON INGALLS) 224 - - - PART VII - - DUST AND FUME RECOVERY - - FLUES, CHAMBERS AND BAG-HOUSES - - DUST CHAMBER DESIGN (MAX J. WELCH) 229 - - CONCRETE IN METALLURGICAL CONSTRUCTION (HENRY W. EDWARDS) 234 - - CONCRETE FLUES (EDWIN H. MESSITER) 240 - - CONCRETE FLUES (FRANCIS T. HAVARD) 242 - - BAG-HOUSES FOR SAVING FUME (WALTER RENTON INGALLS) 244 - - - PART VIII - - BLOWERS AND BLOWING ENGINES - - ROTARY BLOWERS VS. BLOWING ENGINES FOR LEAD SMELTING (EDITORIAL) 251 - - ROTARY BLOWERS VS. BLOWING ENGINES (J. PARKE CHANNING) 254 - - BLOWERS AND BLOWING ENGINES FOR LEAD AND COPPER SMELTING - (HIRAM W. HIXON) 256 - - BLOWING ENGINES AND ROTARY BLOWERS (S. E. BRETHERTON) 258 - - - PART IX - - LEAD REFINING - - THE REFINING OF LEAD BULLION (F. L. PIDDINGTON) 263 - - THE ELECTROLYTIC REFINING OF BASE LEAD BULLION (TITUS ULKE) 270 - - ELECTROLYTIC LEAD REFINING (ANSON G. BETTS) 274 - - - PART X - - SMELTING WORKS AND REFINERIES - - THE NEW SMELTER AT EL PASO, TEXAS (EDITORIAL) 285 - - NEW PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY AT - MURRAY, UTAH (WALTER RENTON INGALLS) 287 - - THE MURRAY SMELTER, UTAH (O. PUFAHL) 291 - - THE PUEBLO LEAD SMELTERS (O. PUFAHL) 294 - - THE PERTH AMBOY PLANT OF THE AMERICAN SMELTING AND REFINING - COMPANY (O. PUFAHL) 296 - - THE NATIONAL PLANT OF THE AMERICAN SMELTING AND REFINING - COMPANY (O. PUFAHL) 299 - - THE EAST HELENA PLANT OF THE AMERICAN SMELTING AND REFINING - COMPANY (O. PUFAHL) 302 - - THE GLOBE PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY - (O. PUFAHL) 304 - - LEAD SMELTING IN SPAIN (HJALMAR ERIKSSON) 306 - - LEAD SMELTING AT MONTEPONI, SARDINIA (ERMINIO FERRARIS) 311 - - - - - PART I - - NOTES ON LEAD MINING - - - - - SOURCES OF LEAD PRODUCTION IN THE UNITED STATES - - BY WALTER RENTON INGALLS - - (November 28, 1903) - - -Statistics of lead production are of value in two directions: (1) in -showing the relative proportion of the kinds of lead produced; and (2) -in showing the sources from which produced. Lead is marketed in three -principal forms: (_a_) desilverized; (_b_) soft; (_c_) antimonial, or -hard. The terms to distinguish between classes “a” and “b” are inexact, -because, of course, desilverized lead is soft lead. Desilverized lead -itself is classified as “corroding,” which is the highest grade, and -ordinary “desilverized.” Soft lead, referring to the Missouri product, -may be either “ordinary” or “chemical hard.” The latter is such lead -as contains a small percentage of copper and antimony as impurities, -which, without making it really hard, increase its resistance against -the action of acids, and therefore render it especially suitable -for the production of sheet to be used in sulphuric-acid chamber -construction and like purposes. The production of chemical hard lead -is a fortuitous matter, depending on the presence of the valuable -impurities in the virgin ores. If present, these impurities go into -the lead, and cannot be completely removed by the simple process of -refining which is practised. Nobody knows just what proportions of -copper and antimony are required to impart the desired property, and -consequently no specifications are made. Some chemical engineers call -for a particular brand, but this is really only a whim, since the same -brand will not be uniformly the same; practically one brand is as good -as another. Corroding lead is the very pure metal, which is suitable -for white lead manufacture. It may be made either from desilverized or -from the ordinary Missouri product; or the latter, if especially pure, -may be classed as corroding without further refining. Antimonial lead -is really an alloy of lead with about 15 to 30 per cent. antimony, -which is produced as a by-product by the desilverizers of base -bullion. The antimony content is variable, it being possible for the -smelter to run the percentage up to 60. Formerly it was the general -custom to make antimonial lead with a content of 10 to 12 per cent. Sb; -later, with 18 to 20 per cent.; while now 25 to 30 per cent. Sb is best -suited to the market. - -The relative values of the various grades of lead fluctuate -considerably, according to the market place, and the demand and supply. -The schedules of the American Smelting and Refining Company make a -regular differential of 10c. per 100 lb. between corroding lead and -desilverized lead in all markets. In the St. Louis market, desilverized -lead used to command a premium of 5c. to 10c. per 100 lb. over ordinary -Missouri; but now they sell on approximately equal terms. Chemical hard -lead sells sometimes at a higher price, sometimes at a lower price, -than ordinary Missouri lead, according to the demand and supply. There -is no regular differential. This is also the case with antimonial -lead.[1] - -The total production of lead from ores mined in the United States in -1901 was 279,922 short tons, of which 211,368 tons were desilverized, -57,898 soft (meaning lead from Missouri and adjacent States) and -10,656 antimonial. These are the statistics of “The Mineral Industry.” -The United States Geological Survey reported substantially the same -quantities. In 1902 the production was 199,615 tons of desilverized, -70,424 tons of soft, and 10,485 tons of antimonial, a total of 280,524 -tons. There is an annual production of 4000 to 5000 tons of white -lead direct from ore at Joplin, Mo., which increases the total lead -production of the United States by, say, 3500 tons per annum. The -production of lead reported as “soft” does not represent the full -output of Missouri and adjacent States, because a good deal of their -ore, itself non-argentiferous, except to the extent of about 1 oz. per -ton in certain districts, is smelted with silver-bearing ores, going -thus into an argentiferous lead; while in one case, at least, the -almost non-argentiferous lead, obtained by smelting the ore unmixed, is -desilverized for the sake of the extra refining. - -Lead-bearing ores are of widespread occurrence in the United States. -Throughout the Rocky Mountains there are numerous districts in which -the ore carries more or less lead in connection with gold and silver. -For this reason, the lead mining industry is not commonly thought of as -having such a concentrated character as copper mining and zinc mining. -It is the fact, however, that upward of 70 per cent. of the lead -produced in the United States is derived from five districts, and in -the three leading districts from a comparatively small number of mines. -The statistics of these for 1901 to 1904 are as follows:[2] - - ┌──────────┬───────────────────────────────┬───────────────────────┬──── - │ │ PRODUCTION, TONS │ PER CENT. │ - │DISTRICT │ 1901 │ 1902 │ 1903 │ 1904 │ 1901│ 1902│ 1903│ 1904│REF. - ├──────────┼───────┼───────┼───────┼───────┼─────┼─────┼─────┼─────┼──── - │Cœur │ │ │ │ │ │ │ │ │ - │d’Alene │ 68,953│ 74,739│ 89,880│ 98,240│ 24.3│ 26.3│ 32.5│ 32.5│_a_ - │Southeast │ │ │ │ │ │ │ │ │ - │Mo. │ 46,657│ 56,550│ 59,660│ 59,104│ 16.4│ 19.9│ 21.2│ 19.6│_b_ - │Leadville,│ │ │ │ │ │ │ │ │ - │Colo. │ 28,180│ 19,725│ 18,177│ 23,590│ 10.0│ 6.9│ 6.6│ 7.8│_c_ - │Park City,│ │ │ │ │ │ │ │ │ - │Utah │ 28,310│ 36,300│ 36,534│ 30,192│ 10.0│ 12.8│ 13.2│ 10.0│_d_ - │Joplin, │ │ │ │ │ │ │ │ │ - │Mo.-Kan. │ 24,500│ 22,130│ 20,000│ 23,600│ 8.6│ 7.8│ 7.2│ 7.8│_e_ - │ ├───────┼───────┼───────┼───────┼─────┼─────┼─────┼─────┤ - │ Total │196,600│209,444│224,251│234,726│ 69.3│ 73.7│ 81.0│ 77.7│ - - - _a._ The production in 1901 and 1902 is computed from direct returns - from the mines, with an allowance of 6 per cent. for loss of lead in - smelting. The production in 1903 and 1904 is estimated at 95 per cent. - of the total lead product of Idaho. - - _b._ This figure includes only the output of the mines at Bonne Terre, - Flat River, Doe Run, Mine la Motte and Fredericktown. It is computed - from the report of the State Lead and Zinc Mine Inspector as to ore - produced, the ore (concentrates) of the mines at Bonne Terre, Flat - River and Doe Run being reckoned as yielding 60 per cent. lead. - - _c._ Report of State Commissioner of Mines. - - _d._ Report of Director of the Mint on “Production of Gold and Silver - in the United States,” with allowance of 6 per cent. for loss of lead - in smelting. - - _e._ From statistics reported by “The Mineral Industry,” reckoning the - ore (concentrates) as yielding 70 per cent. lead. - -Outside of these five districts, the most of the lead produced in the -United States is derived from other camps in Idaho, Colorado, Missouri -and Utah, the combined output of all other States being insignificant. -It is interesting to examine the conditions under which lead is -produced in the five principal districts. - -_Leadville, Colo._—The mines of Leadville, which once were the largest -lead producers of the United States, became comparatively unimportant -after the exhaustion of the deposits of carbonate ore, but have -attained a new importance since the successful introduction of means -for separating the mixed sulphide ore, which occurs there in very large -bodies. The lead production of Leadville in 1897 was 11,850 tons; -17,973 tons in 1898; 24,299 tons in 1899; 31,300 tons in 1900; 28,180 -tons in 1901, and 19,725 tons in 1902. The Leadville mixed sulphide ore -assays about 8 per cent. Pb, 25 per cent. Zn and 10 oz. silver per ton. -It is separated into a zinc product assaying about 38 per cent. Zn and -6 per cent. Pb, and a galena product assaying about 45 per cent. Pb, 10 -or 12 per cent. Zn, and 10 oz. silver per ton. - -_Cœur d’Alene._—The mines of this district are opened on fissure veins -of great extent. The ore is of low grade and requires concentration. As -mined, it contains about 10 per cent. lead and a variable proportion -of silver. It is marketed as mineral, averaging about 50 per cent. Pb -and 30 oz. silver per ton. The production of lead ore in this district -is carried on under the disadvantages of remoteness from the principal -markets for pig lead, high-priced labor, and comparatively expensive -supplies. It enjoys the advantages of large orebodies of comparatively -high grade in lead, and an important silver content, and in many cases -cheap water power, and the ability to work the mines through adit -levels. The cost of mining and milling a ton of crude ore is $2.50 to -$3.50. The mills are situated, generally, at some distance from the -mines, the ore being transported by railway at a cost of 8 to 20c. -per ton. The dressing is done in large mills at a cost of 40 to 50c. -per ton. About 75 per cent. of the lead of the ore is recovered. The -concentrates are sold at 90 per cent. of their lead contents and 95 per -cent. of their silver contents, less a smelting charge of $8 per ton, -and a freight rate of $8 per ton on ore of less than $50 value per ton, -$10 on ore worth $50 to $65, and $12 on ore worth more than $65; the -ore values being computed f. o. b. mines. The settling price of lead is -the arbitrary one made by the American Smelting and Refining Company. -With lead (in ore) at 3.5c. and silver at 50c., the value, f. o. b. -mines, of a ton of ore containing 50 per cent. Pb and 30 oz. silver is -approximately as follows: - - 1000 × 0.90 = 900 lb. lead, at 3.5c. $31.50 - 30 × 0.95 = 28.5 oz. silver, at 50c. 14.25 - —————— - Gross value, f. o. b. mines $45.75 - Less freight, $10, and smelting charge, $8 18.00 - —————— - Net value, f. o. b. mines $27.75 - -Assuming an average of 6 tons of crude ore to 1 ton of concentrate, the -value per ton of crude ore would be about $4.62½, and the net profit -per ton about $1.62½, which figures are increased 23.75c. for each 5c. -rise in the value of silver above 50c. per ounce. - -The production of the Cœur d’Alene since 1895, as reported by the -mines, has been as follows: - - ─—─—─—-———─—─┬—─—-——————─—─┬—─—-—————————┬——————─—─ - YEAR │ LEAD, TONS │ SILVER, OZ. │ RATIO[3] - ─—─—─—-———─—─┼—─—-——————─—─┼—─—-—————————┼——————─—─ - 1896 │ 37,250 │ 2,500,000 │ 67.1 - 1897 │ 57,777 │ 3,579,424 │ 61.9 - 1898 │ 56,339 │ 3,399,524 │ 60.3 - 1899 │ 50,006 │ 2,736,872 │ 54.7 - 1900 │ 81,535 │ 4,755,877 │ 58.3 - 1901 │ 68,953 │ 3,349,533 │ 48.5 - 1902 │ 74,739 │ 4,489,549 │ 60.0 - 1903 │ [4]100,355 │ 5,751,613 │ 57.3 - 1904 │ [4]108,954 │ 6,247,795 │ 57.4 - ─—─—─—-———─—─┴—─—-——————─—─┴—─—-—————————┴——————─—─ - -The number of producers in the Cœur d’Alene district is comparatively -small, and many of them have recently consolidated, under the name of -the Federal Mining and Smelting Company. Outside of that concern are -the Bunker Hill & Sullivan, the Morning and the Hercules mines, control -of which has lately been secured by the American Smelting and Refining -Company. - -_Southeastern Missouri._—The most of the lead produced in this region -comes from what is called the disseminated district, comprising -the mines of Bonne Terre, Flat River, Doe Run, Mine la Motte and -Fredericktown, of which those of Bonne Terre and Flat River are the -most important. The ore of this region is a magnesian limestone -impregnated with galena. The deposits lie nearly flat and are very -large. They yield about 5 per cent. of mineral, which assays about 65 -per cent. lead. The low grade of the ore is the only disadvantage which -this district has, but this is so much more than offset by the numerous -advantages, that mining is conducted very profitably, and it is an open -question whether lead can be produced more cheaply here or in the Cœur -d’Alene. The mines of southeastern Missouri are only 60 to 100 miles -distant from St. Louis, and are in close proximity to the coalfields -of southern Illinois, which afford cheap fuel. The ore lies at depths -of only 100 to 500 ft. below the surface. The ground stands admirably, -without any timbering. Labor and supplies are comparatively cheap. -Mining and milling can be done for $1.05 to $1.25 per ton of crude ore, -when conducted on the large scale that is common in this district. -Most of the mining companies are equipped to smelt their own ore, the -smelters being either at the mines or near St. Louis. The freight rate -on concentrates to St. Louis is $1.40 per ton; on pig lead it is $2.10 -per ton. The total cost of producing pig lead, delivered at St. Louis, -is about 2.25c. per pound, not allowing for interest on the investment, -amortization, etc. - -The production of the mines in the disseminated district in 1901 was -equivalent to 46,657 tons of pig lead; in 1902 it was 56,550 tons. The -milling capacity of the district is about 6000 tons per day, which -corresponds to a capacity for the production of about 57,000 tons of -pig lead per annum. The St. Joseph Lead Company is building a new 1000 -ton mill, and the St. Louis Smelting and Refining Company (National -Lead Company) is further increasing its output, which improvements will -increase the daily milling capacity by about 1400 tons, and will enable -the district to put out upward of 66,000 tons of pig lead. In this -district, as in the Cœur d’Alene, the industry is closely concentrated, -there being only nine producers, all told. - -_Park City, Utah._—Nearly all the lead produced by this camp comes -from the Silver King, Daly West, Ontario, Quincy, Anchor and Daly -mines, which have large veins of low-grade ore carrying argentiferous -galena and blende, a galena product being obtained by dressing, and -zinkiferous tailings, which are accumulated for further treatment as -zinc ore, when market conditions justify.[5] - -_Joplin District._—The lead production of southwestern Missouri and -southeastern Kansas, in what is known as the Joplin district, is -derived entirely as a by-product in dressing the zinc ore of that -district. It is obtained as a product assaying about 77 per cent. Pb, -and is the highest grade of lead ore produced, in large quantity, -anywhere in the United States. It is smelted partly for the production -of pig lead, and partly for the direct manufacture of white lead. The -lead ore production of the district was 31,294 tons in 1895, 26,927 -tons in 1896, 29,578 tons in 1897, 26,457 tons in 1898, 24,100 tons -in 1899, 28,500 tons in 1900, 35,000 tons in 1901, and 31,615 tons in -1902. The production of lead ore in this district varies more or less, -according to the production of zinc ore, and is not likely to increase -materially over the figure attained in 1901. - - - - - NOTES ON THE SOURCE OF THE SOUTHEAST MISSOURI LEAD - - BY H. A. WHEELER - - (March 31, 1904) - - -The source of the lead that is being mined in large quantities in -southeastern Missouri has been a mooted question. Nor is the origin -of the lead a purely theoretical question, as it has an important -bearing on the possible extension of the orebodies into the underlying -sandstone. - -The disseminated lead ores of Missouri occur in a shaly, magnesian -limestone of Cambrian age in St. François, Madison and Washington -counties, from 60 to 130 miles south of St. Louis. The limestone -is known as the Bonne Terre, or lower half of “the third magnesian -limestone” of the Missouri Geological Survey, and rests on a sandstone, -known as “the third sandstone,” that is the base of the sedimentary -formations in the area. Under this sandstone occur the crystalline -porphyries and granites of Algonkian and Archean age, which outcrop as -knobs and islands of limited extent amid the unaltered Cambrian and -Lower Silurian sediments. - -The lead occurs as irregular granules of galena scattered through the -limestone in essentially horizontal bodies that vary from 5 to 100 -ft. in thickness, from 25 to 500 ft. in width, and have exceeded 9000 -ft. in length. There is no vein structure, no crushing or brecciation -of the inclosing rock, yet these orebodies have well defined axes or -courses, and remarkable reliability and persistency. It is true that -the limestone is usually darker, more porous, and more apt to have thin -seams of very dark (organic) shales where it is ore-bearing than in the -surrounding barren ground. The orebodies, however, fade out gradually, -with no sharp line between the pay-rock and the non-paying, and the -lead is rarely, if ever, entirely absent in any extent of the limestone -of the region. While the main course of the orebodies seems to be -intimately connected with the axes of the gentle anticlinal folds, -numerous cross-runs of ore that are associated with slight faults are -almost as important as the main shoots, and have been followed for -5000 ft. in length. These cross-runs are sometimes richer than the -main runs, at least near the intersections, but they are narrower, and -partake more of the type of vertical shoots, as distinguished from the -horizontal sheet-form. - -Most of the orebodies occur at, or close to, the base of the limestone, -and frequently in the transition rock between the underlying sandstone -and the limestone, though some notable and important bodies have -been found from 100 to 200 ft. above the sandstone. This makes the -working depth from the surface vary from 150 to 250 ft., for the upper -orebodies, to 300 to 500 ft. deep to the main or basal orebodies, -according as erosion has removed the ore-bearing limestone. The -thickness of the latter ranges from 400 to 500 ft. - -Associated with the galena are less amounts of pyrite, which especially -fringes the orebodies, and very small quantities of chalcopyrite, zinc -blende, and siegenite (the double sulphide of nickel and cobalt). -Calcite also occurs, especially where recent leaching has opened -vugs, caves, or channels in the limestone, when secondary enrichment -frequently incrusts these openings with crystals of calcite and galena. -No barite ever occurs with the disseminated ore, though it is the -principal gangue mineral in the upper or Potosi member of the third -magnesian limestone, and is never absent in the small ore occurrences -in the still higher second magnesian limestone. - -While the average tenor of the ore is low, the yield being from 3 to -4 per cent. in pig lead, they are so persistent and easy to mine that -the district today is producing about 70,000 tons of pig lead annually, -and at a very satisfactory profit. As the output was about 2500 tons -lead in 1873, approximately 8500 tons in 1883, and about 20,000 tons in -1893, it shows that this district is young, for the principal growth -has been within the last five years. - -Of the numerous but much smaller occurrences of lead elsewhere in -Missouri and the Mississippi valley, none resembles this district -in character, a fact which is unique. For while the Mechernich lead -deposits, in Germany, are disseminated, and of even lower grade than in -Missouri, they occur in a sandstone, and (like all the lead deposits -outside of the Mississippi valley) they are argentiferous, at least to -an extent sufficient to make the extraction of the silver profitable; -and on the non-argentiferous character of the disseminated deposits -hangs my story. - -Of the numerous hypotheses advanced to account for the origin of these -deposits, there are only two that seem worthy of consideration: (a) -the _lateral secretion theory_, and (b) _deposition from solutions of -deep-seated origin_. Other theories evolved in the pioneer period of -economic geology are interesting, chiefly by reason of the difficulties -under which the early strugglers after geological knowledge blazed the -pathway for modern research and observation. - -The lateral secretion theory, as now modernized into the secondary -enrichment hypothesis, has much merit when applied to the southeastern -and central Missouri lead deposits. For the limestones throughout -Missouri—and they are the outcropping formation over more than half of -the State—are rarely, if ever, devoid of at least slight amounts of -lead and zinc, although they range in age from the Carboniferous down -to the Cambrian. - -The sub-Carboniferous formation is almost entirely made up of -limestones, which aggregate 1200 to 1500 ft. in thickness. They -frequently contain enough lead (and less often zinc) to arouse the -hopes of the farmer, and more or less prospecting has been carried on -from Hannibal to St. Louis, or 125 miles along the Mississippi front, -and west to the central part of the State, but with most discouraging -results. - -In the rock quarries of St. Louis, immediately under the lower coal -measures, fine specimens of millerite of world-wide reputation occur -as filiform linings of vugs in this sub-Carboniferous limestone. These -vugs occur in a solid, unaltered rock which gives no clue to the -existence of the vug or cavity until it is accidentally broken. The -vugs are lined with crystals of pink dolomite, calcite and millerite, -with occasionally barite, selenite, galena and blende. They occur -in a well-defined horizon about 5 ft. thick, and the vugs in the -limestone above and below this millerite bed contain only calcite, -or less frequently dolomite. Yet this sub-Carboniferous formation in -southwestern Missouri, about Joplin, carries the innumerable pockets -and sheets of lead and zinc that have made that district the most -important zinc producer in the world. While faulting and limited -folding occur in eastern and central Missouri to fully as great an -extent as in St. François county or the Joplin district, thus far no -mineral concentration into workable orebodies has been found in this -formation, except in the Joplin area. - -The next important series of limestones that make up most of the -central portion of Missouri are of Silurian age, and in them lead and -zinc are liberally scattered over large areas. In the residual surface -clays left by dissolution of the limestone, the farmers frequently make -low wages by gophering after the liberated lead, and the aggregate -of these numerous though insignificant gopher-holes makes quite a -respectable total. But they are only worked when there is nothing else -to do on the farm, as with rare exceptions they do not yield living -wages, and the financial results of mining the rock are even less -satisfactory. Yet a few small orebodies have been found that were -undoubtedly formed by local leaching and re-precipitation of this -diffused lead and zinc. Such orebodies occur in openings or caves, -with well crystallized forms of galena and blende, and invariably -associated with crystallized “tiff” or barite. I am not aware of any -of these pockets or secondary enrichments having produced as much as -2000 tons of lead or zinc, and very few have produced as much as 500 -tons, although one of these pockets was recently exploited with heroic -quantities of printer’s ink as the largest lead mine in the world. Yet -there are large areas in which it is almost impossible to put down a -drill-hole without finding “shines” or trifling amounts of lead or -zinc. That these central Missouri lead deposits are due to lateral -secretion there seems little doubt, and it is possible that larger -pockets may yet be found where more favorable conditions occur. - -When the lateral secretion theory is applied to the disseminated -deposits of southeastern Missouri, we are confronted by enormous bodies -of ore, absence of barite, non-crystallized condition of the galena -except in local, small, evidently secondary deposits, and well-defined -courses for the main and cross-runs of ore. The Bonne Terre orebody, -which has been worked longest and most energetically, has attained a -length of nearly 9000 ft., with a production of about 350,000 tons -or $30,000,000 of lead, and is far from being exhausted. Orebodies -recently opened are quite as promising. The country rock is not as -broken nor as open as in central Missouri, and is therefore much less -favorable for the lateral circulation of mineral waters, yet the -orebodies vastly exceed those of the central region. - -Further, the Bonne Terre formation is heavily intercalated with thick -sheets of shale that would hinder overlying waters from reaching the -base of the ore-horizon, where most of the ore occurs, so that the -leachable area would be confined to a very limited vertical range, -or to but little greater thickness than the 100 ft. or so in which -most of the orebodies occur. While I have always felt that such large -bodies, showing relatively rapid precipitation of the lead, could not -be satisfactorily explained except as having a deep-seated origin, -the fact that the disseminated ore is practically non-argentiferous, -or at least carries only one to three ounces per ton, has been a -formidable obstacle. For the lead in the small fissure-veins that -occasionally occur in the adjacent granite has always been reported -as argentiferous. Thus the Einstein silver mine, near Fredericktown, -worked a fissure-vein from 1 to 6 ft. wide in the granite. It had a -typical complex vein-filling and structure, and carried galena that -assayed from 40 to 200 oz. per ton. While the quantity of ore obtained -did not justify the expensive plant erected to operate it, the galena -was rich in silver, whereas in the disseminated ores at the Mine la -Motte mine, ten miles distant, only the customary 1.5 oz. per ton -occurs. Occasionally fine-grained specimens of galena that I have found -in the disseminated belt would unquestionably be rated as argentiferous -by a Western miner, but the assay showed that the structure in this -case was due to other causes, as only about two ounces were found. -An apparent exception was reported at the Peach Orchard diggings, -in Washington county, in the higher or Potosi member of the third -magnesian limestone, where Arthur Thacher found sulphide and carbonate -ore carrying 8 to 10 oz. of silver per ton; and a short-lived hamlet, -known as Silver City, sprang up to work them. I found, however, that -these deposits are associated with little vertical fissure-veins or -seams that unquestionably come up from the underlying porphyry. - -Recently I examined the Jackson Revel mine, which has been considered -a silver mine for the last fifty years. It lies about seven miles -south of Fredericktown, and is a fissure-vein in Algonkian felsite, -where it protrudes, as a low hill, through the disseminated limestone -formation. A shaft has just been sunk about 150 ft. at less than -1000 ft. from the feather edge of the limestone. The vein is narrow, -only one to twelve inches wide, with slicken-sided walls, runs about -N. 20 deg. E., and dips 80 to 86 deg. eastward. White quartz forms -the principal part of the filling; the vein contains more or less -galena and zinc blende. Assays of the clean galena made by Prof. W. B. -Potter show only 2.5 oz. silver per ton, or no more than is frequently -found in the disseminated lead ores. As the lead in this fissure vein -may be regarded as of undoubted deep origin, and it is practically -non-argentiferous, this would seem to remove the last objection to -the theory of the deep-seated source of the lead in the disseminated -deposits of southeast Missouri. - - - - - MINING IN SOUTHEASTERN MISSOURI - - BY WALTER RENTON INGALLS - - (February 18, 1904) - - -The St. Joseph Lead Company, in the operation of its mines at Bonne -Terre, does not permit the cages employed for hoisting purposes to be -used for access to the mine. Men going to and from their work must -climb the ladders. This rule does not obtain in the other mines of the -district. The St. Joseph Lead Company employs electric haulage for -the transport of ore underground at Bonne Terre. In the other mines -of the district, mules are generally used. The flow of water in the -mines of the district is extremely variable; some have very little; -others have a good deal. The Central mine is one of the wettest in the -entire district, making about 2000 gal. of water per minute. Coal in -southeastern Missouri costs $2 to $2.25 per ton delivered at the mines, -and the cost of raising 2000 gal. of water per minute from a depth -of something like 350 ft. is a very considerable item in the cost of -mining and milling, which, in the aggregate, is expected to come to not -much over $1.25 per ton. - -The ore shoots in the district are unusually large. Their precise trend -has not been identified. Some consider the predominance of trend to be -northeast; others, northwest. They go both ways, and appear to make -the greatest depositions of ore at their intersections. However, the -network of shoots, if that be the actual occurrence, is laid out on a -very grand scale. Vertically there is also a difference. Some shafts -penetrate only one stratum of ore; others, two or three. The orebody -may be only a few feet in thickness; it may be 100 ft. or more. The -occurrence of several overlying orebodies obviously indicates the -mineralization of different strata of limestone, while in the very -thick orebodies the whole zone has apparently been mineralized. - -The grade of the ore is extremely variable. It may be only 1 or 2 per -cent. mineral, or it may be 15 per cent. or more. However, the average -yield for the district, in large mines which mill 500 to 1200 tons of -ore per day, is probably about 5 per cent. of mineral, assaying 65 per -cent. Pb, which would correspond to a yield of 3.25 per cent. metallic -lead in the form of concentrate. The actual recovery in the dressing -works is probably about 75 per cent., which would indicate a tenure of -about 4.33 per cent. lead in the crude ore. - - - - - LEAD MINING IN SOUTHEASTERN MISSOURI - - BY R. D. O. JOHNSON - - (September 16, 1905) - - -The lead deposits of southeastern Missouri carry galena disseminated -in certain strata of magnesian limestone. Their greater dimensions -are generally horizontal, but with outlines extremely irregular. The -large orebodies consist usually of a series of smaller bodies disposed -parallel to one another. These smaller members may coalesce, but are -generally separated from one another by a varying thickness of lean ore -or barren rock. The vertical and lateral dimensions of an orebody may -be determined with a fair degree of accuracy by diamond drilling, and a -map may be constructed from the information so obtained. Such a map (on -which are plotted the surface contours) makes it possible to determine -closely the proper location of the shaft, or shafts, considering also -the surface and underground drainage and tramming. - -The first shafts in the district were sunk at Bonne Terre, where the -deposits lie comparatively near the surface. The early practice at this -point was to sink a number of small one-compartment shafts. As the -deposits were followed deeper, this gave way to the practice of putting -down two-compartment shafts equipped much more completely than were the -shallower shafts. - -At Flat River (where the deposits lie at much greater depths, some -being over 500 ft.) the shafts are 7 × 14 ft., 6½ × 18 ft., and 7 × 20 -ft. These larger dimensions give room not only for two cage-ways and a -ladder-way, but also for a roomy pipe-compartment. The large quantities -of water to be pumped in this part of the district make the care of -the pipes in the shafts a matter of first importance. At Bonne Terre -only such a quantity of water was encountered as could be handled by -bailing or be taken out with the rock; there the only pipe necessary -was a small air-pipe down one corner of the shaft. When the quantity -of water encountered is so great that the continued working of the mine -depends upon its uninterrupted removal, the care of the pipes becomes a -matter of great importance. A shaft which yields from 4000 to 5000 gal. -of water per minute is equipped with two 12 in. column pipes and two 4 -in. steam pipes covered and sheathed. Moreover, the pipe compartment -will probably contain an 8 in. air-pipe, besides speaking-tubes, pipes -for carrying electric wires, and pipes for conducting water from upper -levels to the sump. To care for these properly there are required a -separate compartment and plenty of room. - -Shafts are sunk by using temporary head frames and iron buckets of from -8 to 14 cu. ft. capacity. Where the influx of water was small, 104 ft. -have been sunk in 30 days, with three 8 hour shifts, two drills, and -two men to each drill; 2¾ in. drills are used almost exclusively; 3¼ -in. drills have been used in sinking, but without apparent increase in -speed. - -The influence of the quantity of water encountered upon the speed of -sinking (and the consequent cost per foot) is so great that figures are -of little value. Conditions are not at all uniform. - -At some point (usually before 200 ft. is reached) a horizontal opening -will be encountered. This opening invariably yields water, the amount -following closely the surface precipitation. It is the practice to -establish at this point a pumping station. The shaft is “ringed” and -the water is directed into a sump on the side, from which it is pumped -out. This sump receives also the discharge of the sinking pumps. - -The shafts sunk in solid limestone require no timbering other than that -necessary to support the guides, pipes, and ladder platforms. These -timbers are 8 × 8 in. and 6 x 8 in., spaced 7 or 8 ft. apart. - -Shafts are sunk to a depth of 10 ft. below the point determined upon -as the lower cage landing. From the end at the bottom a narrow drift -is driven horizontally to a distance of 15 ft.; at that point it is -widened out to 10 ft. and driven 20 ft. further. The whole area (10 × -20 ft.) is then raised to a point 28 or 30 ft. above the bottom of the -drift from the shaft. The lower part of this chamber constitutes the -sump. Starting from this chamber (on one side and at a point 10 ft. -above the cage landing, or 20 ft. above the bottom of the sump), the -“pump-house” is cut out. This pump-house is cut 40 ft. long and is as -wide as the sump is long, namely, 20 ft. A narrow drift is driven to -connect the top of the pump-house with the shaft. Through this drift -the various pipes enter the pump-house from the shaft. - -The pumps are thus placed at an elevation of 10 ft. above the bottom -of the mine. Flooding of mines, due to failure of pumps or to striking -underground bodies of water, taught the necessity of placing the pumps -at such an elevation that they would be the last to be covered, thus -giving time for getting or keeping them in operation. The pumps are -placed on the solid rock, the air pumps and condensers at a lower level -on timbers over the sump. - -With this arrangement, the bottom of the shaft serves as an antechamber -for the sump, in which is collected the washing from the mine and the -dripping from the shaft. The sump proper rarely needs cleaning. - -The pumps are generally of high-grade, compound-and triple-expansion, -pot-valved, outside-packed plunger pattern. Plants with electrical -power distribution have recently installed direct-connected compound -centrifugal pumps with entire success. - -Pumps of the Cornish pattern have never been used much in this region. -One such pump has been installed, but the example has not been followed -even by the company putting it in. - -The irregular disposition of the ore renders any systematic plan of -drifting or mining (as in coal or vein mining) impossible. On each -side of the shaft and in a direction at right angles to its greater -horizontal dimension, drifts 12 to 14 ft. in width are driven to a -distance of 60 or 70 ft. In these broad drifts are located the tracks -and also the “crossovers” for running the cars on and off the cage. - -When a deposit is first opened up, it is usually worked on two, and -sometimes three, levels. These eventually cut into one another, when -the ore is hoisted from the lower level alone. - -The determination of the depth of the lower level is a matter of -compromise. Much good ore may be known to exist below; when it comes -to mining, it will have to be taken out at greater expense; but the -level is aimed to cut through the lower portions of the main body. It -is generally safe to predict that the ore lying below the upper levels -will eventually be mined from a lower level without the expense of -local underground hoisting and pumping. - -The ore has simply to be followed; no one can say in advance how it -is going to turn out. The irregularity of the deposits renders any -general plan of mining of little or no value. Some managers endeavor to -outline the deposits by working on the outskirts, leaving the interior -as “ore reserves.” Such reserves have been found to be no reserves at -all, though the quality of the rock may be fairly well determined by -underground diamond drilling. Many of the deposits are too narrow to -permit the employment of any system of outlining and at the same time -keeping up the ore supply. - -The individual bodies constituting the general orebody are rarely, -if ever, completely separated by barren rock; some “stringers” or -“leaders” of ore connect them. The careful superintendent keeps a -record on the monthly mine map of all such occurrences, or otherwise, -or of blank walls of barren rock that mark the edge of the deposit. -This precaution finds abundant reward when the drills commence to “cut -poor,” and when a search for ore is necessary. - -The method of mining is to rise to the top of the ore and to carry -forward a 6 ft. breast. If the ore is thick enough, this is followed by -the underhand stope. Drill holes in the breast are usually 7 or 8 ft. -in depth; stope holes, 10 to 14 feet. - -Both the roof and the floor are drilled with 8 or 10 ft. holes placed -8 or 10 ft. apart. These serve to prospect the rock in the immediate -neighborhood; in the roof they serve the further very important purpose -of draining out water that might otherwise accumulate between the -strata and that might force them to fall. The condition or safety -of the roof is determined by striking with a hammer. If the sound -is hollow or “drummy,” the roof is unsafe. If water is allowed to -accumulate between the loose strata, obviously it is not possible to -determine the condition of the roof. - -It is the duty of two men on each shift to keep the mine in a safe -condition by taking down all loose and dangerous masses of rock. These -men are known as “miners.” It sometimes happens that a considerable -area of the roof gets into such a dangerous condition that it is either -too risky or too expensive to put in order, in which case the space -underneath is fenced off. As a general thing, the mines are safe and -are kept so. There are but few accidents of a serious nature due to -falling rock. - -The roof is supported entirely by pillars; no timbering whatever is -used. The pillars are parts of the orebody or rock that is left. They -are of all varieties of size and shape. They are usually circular in -cross-section, 10 to 15 ft. in diameter and spaced 20 to 35 ft. apart, -depending upon the character of the roof. Pillars generally flare at -the top to give as much support to the roof as possible. The hight of -the pillars corresponds, of course, to the thickness of the orebody. - -All drilling is done by 2¾ in. percussion drills. In the early days, -when diamonds were worth $6 per carat, underground diamond drills were -used. Diamond drills are used now occasionally for putting in long -horizontal holes for shooting down “drummy” roof. Air pressure varies -from 60 to 80 lb. Pressures of 100 lb. and more have been used, but the -repairs on the drills became so great that the advantages of the higher -pressure were neutralized. - -Each drill is operated by two men, designated as “drillers,” or “front -hand” and “back hand.” The average amount of drilling per shift of 10 -hours is in the neighborhood of 40 ft., though at one mine an average -of 55 ft. was maintained. - -In some of the mines the “drillers” and “back hands” do the loading and -firing; in others that is done by “firers,” who do the blasting between -shifts. When the drillers do the firing, there is employed a “powder -monkey,” who makes up the “niphters,” or sticks of powder, in which are -inserted and fastened the caps and fuse; 35 per cent. powder is used -for general mining. - -Battery firing is employed only in shaft sinking. In the mining work -this is found to be much more expensive; the heavy concussions loosen -the stratum of the roof and make it dangerous. - -Large quantities of oil are used for lubrication and illumination. -“Zero” black oil and oils of that grade are used on the drills. Miners’ -oil is generally used for illumination, though some of the mines use a -low grade of felsite wax. - -Two oil cans (each holding about 1½ pints) are given to each pair of -drillers, one can for black oil and one for miners’ oil. These cans, -properly filled, are given out to the men, as they go on shift, at the -“oil-house,” located near the shaft underground. This “oil-house” is in -charge of the “oil boy,” whose duty it is to keep the cans clean, to -fill them and to give them out at the beginning of the shift. The cans -are returned to the oil-house at the end of the shift. - -Kerosene is used in the hat-lamps in wet places. - -The “oil-houses” are provided with three tanks. In some instances these -tanks are charged through pipes coming down the shaft from the surface -oil-house. These tanks are provided with oil-pumps and graduated -gage-glasses. - -Shovelers or loaders operate in gangs of 8 to 12, and are supervised by -a “straw boss,” who is provided with a gallon can for illuminating oil. -The cars are 20 cu. ft. (1 ton) capacity. Under ordinary conditions one -shoveler will load 20 of these cars in a shift of 10 hours. They use -“half-spring,” long-handled, round-pointed shovels. - -Cars are of the solid-box pattern, and are dumped in cradles. Loose -and “Anaconda” manganese-steel wheels are the most common. Gage of -track, 24 to 30 in., 16 lb. rails on main lines and 12 lb. on the side -and temporary tracks. Cars are drawn by mules. One mine has installed -compressed-air locomotives and is operating them with success. - -Shafts are generally equipped with geared hoists, both steam and -electrically driven. Later hoists are all of the first-motion pattern. - -Generally the cars are hoisted to the top and dumped with cradles. One -shaft, however, is provided with a 5-ton skip, charged at the bottom -from a bin, into which the underground cars are dumped. Upon arriving -at the top the skip dumps automatically. This design exhibits a number -of advantages over the older method and will probably find favor with -other mine operators. - - - - - THE LEAD ORES OF SOUTHWESTERN MISSOURI - - BY C. V. PETRAEUS AND W. GEO. WARING - - (October 21, 1905) - - -The lead ore of southwestern Missouri, and the adjoining area in the -vicinity of Galena, Kan., is obtained as a by-product of zinc mining, -the galena being separated from the blende in the jigging process. -Formerly the galena (together with “dry-bone,” including cerussite and -anglesite) was the principal ore mined from surface deposits in clay, -the blende being the subsidiary product. In the deeper workings blende -was found largely to predominate; this is shown by the shipments of the -district in 1904, which amounted to 267,297 tons of zinc concentrate -and 34,533 tons of lead concentrate. - -The lead occurs in segregated cubes, from less than one millimeter up -to one foot in diameter. The cleavage is perfect, so that each piece -of ore when struck with a hammer breaks up into smaller perfect cubes. -In this respect the ore differs from the galena encountered in the -Rocky Mountain regions, where torsional or shearing strains seem in -most instances to have destroyed the perfect cleavage of the minerals -subsequent to their original deposition. Cases of schistose and twisted -structure occur in lead deposits of the Joplin district but rarely, and -they are always quite local. - -The separation of the galena from the blende and marcasite (“mundic”) -in the ordinary process of jigging is very complete; the percentage -of zinc and iron in the lead concentrate is insignificant. As an -illustration of this, the assays of 100 recent consecutive shipments of -lead ore from the district, taken at random, are cited as follows: - - 7 shipments assayed from 57 to 70% lead - 15 shipments assayed from 70 to 75% lead - 46 shipments assayed from 75 to 79% lead - 32 shipments assayed from 80 to 84.4% lead - Average of 100 shipments 78.4% lead - -Fourteen shipment samples, ranging from 70 to 84.4 per cent. lead, were -tested for zinc and iron. These averaged 2.24 per cent. Fe and 1.78 -per cent. Zn, the highest zinc content being 4.5 per cent. No bismuth -or arsenic, and only very minute traces of antimony, have ever been -found in these ores. They contain only about 0.0005 per cent. of silver -(one-seventh of an ounce per ton) and scarcely more than that of copper -(occurring as chalcopyrite). - -The pig lead produced from these ores is therefore very pure, soft and -uniform in quality, so that the term “soft Missouri lead” has become a -synonym for excellence in the manufacture of lead alloys and products, -such as litharge, red and white lead, and orange mineral. Its freedom -from bismuth, which is generally present in Colorado lead, makes it -particularly suitable for white lead; also for glass-maker’s litharge -and red lead. These oxides, for use in making crystal glass, must be -made by double refining so as to remove even the small quantities -of silver and copper that are present. The resulting product, made -from soft Missouri lead, is far superior to any refined lead produced -anywhere in this country or in Europe, even excelling the famous -Tarnowitz lead. It gives a luster and clarity to the glass that no -other lead will produce. Lead from southeastern Missouri, Kentucky, -Illinois, Iowa, and Wisconsin yields identical results, but the -refining is more difficult, not only because the lead contains a little -more silver and copper, but also because it contains more antimony. - -The valuation of the lead concentrate produced in the Joplin district -is based upon a wet assay, usually the molybdate or ferrocyanide -method. The price paid is determined variously. One buyer pays a -fixed price for average ore, making no deductions; as, for example, -at present rates, $32.25 per 1000 lb. whether the ore assays 75 or 84 -per cent. Pb, pig lead being worth $4.75 at St. Louis.[6] Another pays -$32.25 for 80 per cent. ore, or over, deducting 50c. per unit for ores -assaying under 80 per cent. Another pays for 90 per cent. of the lead -content of the ore as shown by the assay, at the St. Louis price of pig -lead, less a smelting charge of, say, $6 to $8 per ton of ore. - -The history of the development of lead ore buying in the Joplin -district is rather curious. In the early days of the district the ore -was smelted wholly on Scotch hearths, which, with the purest ores, -would yield 70 per cent. metallic lead. No account was taken of the -lead in the rich slag, chemical determinations being something unknown -in the district at that time; it being supposed generally that pure -galena contained 700 lb. lead to the 1000 lb. of ore, the value of -700 lb. lead, less $4.50 per 1000 lb. of ore for freight and smelting -costs, was returned to the miner. The buyers graded the ore, according -to their judgment, by its appearance, as to its purity and also as to -its behavior in smelting; an ore, for example, from near the surface, -imbedded in the clay and coated more or less with sulphate, yielded its -metal more freely than the purer galenas from deeper workings. - -This was the origin of the present method of buying—a system that would -hardly be tolerated except for the fact that the lead is, as previously -stated, considered a by-product of zinc mining. - -Originally all the lead ore from the Missouri-Kansas district was -smelted in the same region, either in the air furnace (reverberatory -sweating-furnace) or in the water-back Scotch hearth. Competition -gradually developed in the market. Lead refiners found the pure -sulphide of special value in the production of oxidized products. -Some of the ore found its way to St. Louis, and even as far away as -Colorado, where it was used to collect silver. Since the formation of -the American Smelting and Refining Company and the greatly increased -output of the immense deposits of lead ore in Idaho, no Missouri lead -ore has gone to Colorado. - -Up to 1901, one concern had more or less the control of the -southwestern Missouri ores. At the present time, lead ore is bought -for smelters in Joplin, Carterville, and Granby, Mo., Galena, Kan., -and Collinsville, Ill., and complaint is heard that present prices are -really too high for the comfort of the smelters. Yet the old principle -of paying for lead ores upon the supposed yield of 70 per cent., -irrespective of the real lead content, is still largely in vogue. - -Any one interested in the matter will find it quite instructive to -calculate the returning charges, or gross profits, in smelting these -ores, on the basis of 70 per cent. recovery, at $32.25 per 1000 lb. of -ore, less 50c. per ton haulage, with lead at $4.77 per 100 lb. at St. -Louis. No deduction, it should be remarked, is ever made for moisture -in lead ores in this district. It is of interest to observe that -Dr. Isaac A. Hourwich estimates (in the U. S. Census Special Report -on Mines and Quarries recently issued) the average lead contents of -the soft lead ores of Missouri in 1902 at 68.2 per cent., taking as a -basis the returns from five leading mining and smelting companies of -Missouri, which reported a product of 70,491 tons of lead from 103,428 -tons of their own and purchased ore. The average prices for lead ore -in 1902 were reported as follows, per 1000 lb.: Illinois, $19.53; -Iowa, $24.48; Kansas, $23.51; Missouri, $22.17; Wisconsin, $23.29; -Rocky Mountain and Atlantic Coast States, $10.90. In 1903, according -to Ingalls (“The Mineral Industry,” Vol. XII), the ore from the Joplin -district commanded an average price of $53 per 2000 lb., while the -average in the southeastern district was $46.81. - - - - - PART II - - ROAST-REACTION SMELTING - - SCOTCH HEARTHS AND REVERBERATORY FURNACES - - - - - LEAD SMELTING IN THE SCOTCH HEARTH - - BY KENNETH W. M. MIDDLETON - - (July 6, 1905) - - -In view of the fact that the Scotch hearth in its improved form is now -coming to the front again to some extent in lead smelting, it may prove -interesting to give a brief account of its present use in the north of -England. - -Admitting that, where preliminary roasting is necessary, the best -results can be obtained with the water-jacketed blast furnace (this -being more especially the case where labor is an expensive item), we -have still as an alternative the method of smelting raw in the Scotch -hearth. At one works, which I recently visited, all the ore was smelted -raw; at another, all the ore received a preliminary roast, and it is -instructive to compare the results obtained in the two cases. The -following data refer to a fairly “free-smelting” galena assaying nearly -80 per cent. of lead. - -When smelting raw ore in the hearth, fully 7½ long tons can be treated -in 24 hours, the amount of lead produced direct from the furnace in the -first fire being 8400 to 9000 lb.; this is equivalent to 56 to 60 per -cent. of lead, the remaining 24 to 20 per cent. going into the fume and -the slag. - -When smelting ore which has received a preliminary roast of two hours, -12,000 lb. of lead is produced direct from the hearth, this being -equivalent to 65 per cent. of the ore. When the ore is roasted, the -output of the hearth is practically the same for all ores of equal -richness; but when smelting raw, if the galena is finely divided, the -output may fall much below that given herewith; while, on the other -hand, under the most favorable conditions it may rise to 12,000 lb. in -24 hours, or even more. - -I had an opportunity of seeing a parcel of galena carrying 84 per cent. -of lead (but broken down very fine) smelted raw. The ore was kept damp -and the blast fairly low; but, in spite of that, a quantity of the ore -was blown into the flue, and only 5100 lb. of lead was produced from -the hearth in 24 hours. - -Galena carrying only 65 per cent. of lead does not give nearly as -satisfactory results when smelted raw in the hearth; barely six tons of -ore can be smelted in 24 hours, and only 4500 to 5400 lb. of lead can -be produced directly. This is equivalent to, say, 43 per cent. of the -ore in the first fire; the remaining 22 per cent. goes into the slag or -to the flue as fume. Moreover, the 65 per cent. ore requires 1500 lb. -of coal in 24 hours, while the 80 per cent. galena uses only 1000 lb. - -Turning now for a moment to the costs of smelting raw and of smelting -after a preliminary roast, we find that (in the case of the two works -we have been considering) the results are all in favor of smelting raw, -so far as a galena carrying nearly 80 per cent. is concerned. - -The cost of smelting, per ton of lead produced, is given herewith: - - -ORE SMELTED RAW - - Smelters’ wages $2.04 - “ coal (425 lb.) 0.38 - ——- - Total $2.42 - -A very small quantity of lime is also used in this case for some ores, -but its cost would never amount to more than 4c. per ton of lead -produced. - - -ORE RECEIVING A PRELIMINARY ROAST - - Roasters’ wages $0.61 - “ coal (425 lb.) 0.65 - Smelters’ wages 1.08 - “ coal (75 lb.) 0.11 - Peat and lime 0.08 - ——- - Total $2.53 - -It should be noted also that the smelters at the works where the ore -was not roasted receive higher pay. In the eight-hour shift they -produce about 1½ tons of lead; and as there are two of them to a -furnace, they make $3.06 between them, or $1.53 each. The two men -smelting roasted ore produce about two tons in an eight-hour shift, and -therefore each receives $1.08 per shift. - -Coming now to fume-smelting in the hearth, we can again compare the -results obtained in smelting raw and after roasting. It is well to -bear in mind, also, that, while only 6½ per cent. of the lead goes in -the fume when smelting roasted ores in the hearth, a considerably -larger proportion is thus lost when smelting raw ores. When fume is -smelted raw, it is best dealt with when containing about 40 per cent. -of moisture. One man attends to the hearth (instead of two as when -smelting ore), and in 24 hours 3000 lb. of lead is produced, the amount -of coal used being 2100 lb. No lime is required. - -When smelting roasted fume, two men attend to the hearth and the output -is 6000 lb. in 24 hours, the amount of coal used being 1800 lb. In this -latter case fluorspar happens to be available (practically free of -cost), and a little of it is used with advantage in fume-smelting, as -well as a small quantity of lime. - -The cost of fume-smelting per ton of lead produced is given herewith: - - -FUME SMELTED RAW - - Smelters’ wages $2.88 - “ coal (1400 lb.) 2.13 - ——-—- - Total $5.01 - - -FUME RECEIVING A PRELIMINARY ROAST - - Roasters’ wages $2.08 - “ coal (1450 lb.) 2.18 - Smelters’ wages 2.04 - “ coal (600 lb.) 0.92 - Peat and lime 0.08 - ———-- - Total $7.30 - -In this case, as in that of ore, the smelter of the raw fume gets -better pay; he has $1.44 per eight-hour shift, while the smelter of the -roasted ore has only $1.02 per eight-hour shift. - -Fume takes four hours to roast, as compared to the two hours taken by -ore. - -From these facts regarding Scotch-hearth smelting, it would seem that -with galena carrying, say, over 70 per cent. lead (but more especially -with ore up to 80 per cent. in lead, and, moreover, fairly free from -impurities detrimental to “free” smelting), very satisfactory results -can be obtained by smelting raw. Against this, however, it must be said -that at the works where the ore is roasted attempts at smelting raw -have been made several times without sufficient success to justify the -adoption of this method, although the ores smelted average 75 per cent. -lead and seem quite suitable for the purpose. - -Probably this may be accounted for by the fact that the method of -running the furnace when raw ore is being smelted is rather different -from that adopted when dealing with roasted ore. Moreover, at the works -under notice the furnaces are not of the most modern construction; and, -as the old custom of dropping a peat in front of the blast every time -the fire is made up still survives, it is necessary to shut off the -blast while this is being done, and the fire is then apt to get rather -slack. - -The gray slag produced in the hearth is smelted in a small blast -furnace, a little poor fume, and sometimes a small quantity of -fluorspar, being added to facilitate the process. Some figures -regarding slag-smelting may be of interest. The slag-smelters produce -9000 lb. of lead in 24 hours. The cost of slag-smelting per ton of lead -produced is as follows: - - Smelters’ wages $1.60 - Coke (1500 lb.) 3.42 - Peat 0.06 - ———-- - Total $5.08 - -Recent analyses of Weardale (Durham county) lead smelted in the Scotch -hearth, and slag-lead smelted in the blast furnace, are given herewith: - - ─────────┬───────────────────┬────────────────────┬────────────────── - │ FUME-LEAD FROM │ SILVER-LEAD FROM │ SLAG-LEAD FROM - │ HEARTH │ HEARTH │ BLAST FURNACE - ─────────┼───────────────────┼────────────────────┼────────────────── - Lead │ 99.957 │ 99.957 │ 99.013 - Silver │ 0.0035 │ 0.0200 │ 0.0142 - │ (1oz. 2dwt. 21gr. │ (6oz. 10dwt. 16gr. │(4oz. 12dwt. 18gr. - │ per Long Ton) │ per Long Ton) │ per Long Ton) - Tin │ nil │ nil │ nil - Antimony │ nil │ nil │ 0.874 - Copper │ nil │ nil │ 0.024 - Iron │ 0.019 │ 0.019 │ 0.023 - Zinc │ nil │ nil │ nil - │ ──────── │ ──────── │ ──────── - │ 99.9795 │ 99.9960 │ 99.9482 - ─────────┴───────────────────┴────────────────────┴────────────────── - -The ordinary form of the Scotch hearth is probably too well known -to need much description. The dimensions which have been found most -suitable are as follows: Front to back, 21 in.; width, 27 in.; depth -of hearth, 8 to 12 in. Formerly the distance from front to back was 24 -in., but this was found too much for the blast and for the men. - -The cast-iron hearth which holds the molten lead is set in brickwork; -if 8 in. deep and capable of holding about ¾ ton of lead, it is quite -large enough. The workstone or inclined plate in front of the hearth -is cast in one piece with it, and has a raised holder on either side -at the lower edge, and a gutter to convey the overflowing lead to the -melting-pot. The latter is best made with a partition and an opening -at the bottom through which clean lead can run, so that it can be -ladled into molds without the necessity for skimming the dross off -the surface. It is well also to have a small fireplace below the -melting-pot. - -On each side of the hearth, and resting on it, is a heavy cast-iron -block, 9 in. thick, 15 in. high, 27 to 28 in. long. To save metal, -these are now cast hollow and air is caused to pass through them. On -the back of the hearth stands another cast-iron block known as the -“pipestone,” through which the blast comes into the furnace. In the -older forms of pipestone the blast comes in through a simple round or -oval pipe, a common size being 3 or 4 in. wide by 2½ in. high, and the -pipestone is not water-cooled. With this construction the hearth will -not run satisfactorily unless the pipestone is set with the greatest -care, so as to have the tuyere exactly in the center, and as there -is no water-cooling the metal quickly burns away when fume is being -smelted. Moreover, the blast is apt to be stopped by slag adhering to -the end of the pipe. As already mentioned, a peat is dropped in front -of the blast every time the fire is made up, with the object of keeping -a clear passage open for the blast. This old custom has, however, -several serious disadvantages; first, it prevents the blast being kept -on continuously; and, second, it makes it necessary to have the hearth -open at the top so that the smelter-man can go in by the side of it. In -this case the ore is fed from the side by the smelter-man, who works -under the large hood placed above the furnace to carry away the fume. -Even when he is engaged in shoveling back the fire from the front and -is not underneath the hood, it is impossible to prevent some fume from -blowing out; and there is much more liability to lead-poisoning than -when the hearth is closed at the top by the chimney and the smelter-men -work from the front. The best arrangement is to have the hearth -entirely closed in by the chimney, except for the opening at the front, -and to have a small auxiliary flue above the workstone leading direct -to the open air to catch any fume that may blow out past the shutter in -front of the hearth. - -In an improved form of pipestone, a pipe connected to the blast-main -fits into the semicircular opening at the back and is driven tight -against a ridge in the flat side of the opening. Going through the -pipestone, the arch becomes gradually flatter, and the blast emerges -into the hearth, about 2 in. above the level of the molten lead, -through an oblong slit 12 in. long by 1 in. wide, with a ledge -projecting 1½ in. immediately above it. The back and front are similar, -so that when one side gets damaged the pipestone can be turned back to -front. - -Water is conveyed in a 2½ in. iron pipe to the pipestone, and after -passing through it is led away from the other end to a water-box, which -stands beside the hearth and into which the red-hot lumps of slag are -thrown to safeguard the smelters from the noxious fumes. - -On the top of the pipestone rests an upper backstone, also of cast -iron; it extends somewhat higher than the blocks at the sides. All this -metal above the level of the lead is necessary because the partially -fused lumps which stick to it have to be knocked off with a long bar, -so that if fire-bricks were used in place of cast iron they would soon -be broken up and destroyed. - -With a covered-in hearth, when the ore is charged from the front, -the following is the method adopted in smelting raw ore: The charge -floats on the molten lead in the hearth, and at short intervals the -two smelters running the furnace ease it up with long bars, which they -insert underneath in the lead. Any pieces of slag adhering to the sides -and pipestone are broken off. After easing up the fire, the lumps of -partially reduced ore, mixed with cinders and slag, are shoveled on -to the back of the fire; the slag is drawn out upon the workstone -(any pieces of ore adhering to it being broken off and returned to -the hearth), and it is then quenched in a water-box placed alongside -the workstone. One or two shovelfuls of coal, broken fairly small -and generally kept damp, are thrown on the fire, together with the -necessary amount of ore, which is also kept damp if in a fine state -of division. It is part of the duty of the two smelters to ladle out -the lead from the melting-pot into the molds. In smelting ore a fairly -strong, steady blast is required, and it is made to blow right through -so as to keep the front of the fire bright. A little lime is thrown on -the front of the fire when the slag gets too greasy. - -When smelting raw fume one man attends to the furnace. It does not -have to be made up nearly as frequently, the work being easier for -one man than smelting ore is for two. The unreduced clinkers and slag -are dealt with exactly as in smelting ore; and coal is also, in this -case, thrown on the back of the fire, but the blast does not blow -right through to the front. On the contrary, the front of the fire is -kept tamped up with fume, which should be of the coherency of a thick -mud. The blast is not so strong as that necessary for ore. The idea is -partially to bake the fume before submitting it to the hottest part of -the furnace, or to the part where the blast is most strongly felt. It -is only when smelting fume that it is necessary to keep the pipestone -water-cooled. - -To start a furnace takes from two to three hours. The hearth is left -full of lead, and this has to be melted before the hearth is in normal -working order. Drawing the fire takes about three-quarters of an hour; -the clinkers are taken off and kept for starting the next run, and the -sides and back of the hearth are cleaned down. - - - - - THE FEDERAL SMELTING WORKS, NEAR ALTON, ILL.[7] - - BY O. PUFAHL - - (June 2, 1906) - - -The works of the Federal Lead Company, near Alton, Ill., were erected -in 1902. They have a connection with the Chicago, Peoria & St. Louis -Railway, by which they receive all their raw materials, and by which -all the lead produced is shipped. - -The ore smelted is galena, with dolomitic gangue, and a small quantity -of pyrites (containing a little copper, nickel, and cobalt) from -southeastern Missouri, and consists chiefly of fine concentrates, -containing 60 to 70 per cent. lead. In addition thereto a small -proportion of lump ore is also smelted. - -A striking feature at these works is the excellent facility for -handling the materials. The bins for the ore, coke and coal are made -of concrete and steel and are filled from cars running on tracks -laid above them. For transporting the materials about the works a -narrow-gage railway with electric locomotives is used. - -The ores are smelted by the Scotch-hearth process. There are 20 hearths -arranged in a row in a building constructed wholly of steel and stone. -The sump (4 × 2 × 1 ft.) of each furnace contains about one ton of -lead. The furnaces are operated with low-pressure blast from a main -which passes along the whole row. The blast enters the furnace from a -wind chest at the back through eight 1 in. iron pipes, 2 in. above the -bath of lead. The two sides and the rear wall are cooled by a cast-iron -water jacket of 1 in. internal width. - -Two men work, in eight-hour shifts, at each of the furnaces, receiving -4.75 and 4.25c. respectively for every 100 lb. of lead produced. The -ore is weighed out and heaped up in front of the furnaces; on the -track near by the coke is wheeled up in a flat iron car with two -compartments. The furnacemen are chiefly negroes. At the side of each -furnace is a small stock of coal, which is used chiefly for maintaining -a small fire under the lead kettle. Only small quantities of coal are -added from time to time during the smelting operation. - -Over each furnace is placed an iron hood, through which the fumes and -gases escape. They pass first through a collecting pipe, extending -through the whole works, to a 1500 ft. dust flue, measuring 10 × 10 -ft., in internal cross-section. Near the middle of this is placed a -fan of 100,000 cu. ft. capacity per minute, which forces the fumes and -gases into the bag-house, where they are filtered through 1500 sacks of -loosely woven cotton cloth, each 25 ft. long and 18 in. in diameter, -and thence pass up a 150 ft. stack. - -The dust recovered in the collecting flue is burnt, together with the -fume caught by the bags, the coal which it contains furnishing the -combustible. It burns smolderingly and frits together somewhat. The -product (chiefly lead sulphate) is then smelted in a shaft furnace, -together with the gray slag from the hearth furnaces. The total -extraction of lead is about 98 per cent., i.e., the combined process -of Scotch-hearth and blast-furnace smelting yields 98 per cent. of the -lead contained in the crude ore. - -The direct yield of lead from the Scotch hearths is about 70 per cent. -They also produce gray slag, containing much lead, which amounts to -about 25 per cent. of the weight of the ore. About equal proportions -of lead pass into the slag and into the flue dust. When working to -the full capacity, with rich ore (80 per cent. lead and more) the 20 -furnaces can produce about 200 tons of lead in 24 hours. The coke -consumption in the hearth furnaces amounts to only 8 per cent. of the -ore. The lead from these furnaces is refined for 30 minutes to one -hour by steam in a cast-iron kettle of 35 tons capacity, and is cast -into bars either alone or mixed with lead from the shaft furnace. The -“Federal Brand” carries nearly 99.9 per cent. lead, 0.05 to 0.1 per -cent. copper, and traces of nickel and cobalt. - -The working up of the between products from the hearth-furnaces is -carried out as follows: Slag, burnt flue dust and roasted matte from -a previous run, together with a liberal proportion of iron slag (from -the iron works at Alton), are smelted in a 12-tuyere blast furnace -for work-lead and matte. The furnace is provided with a lead well at -the back. The matte and slag are tapped off together at the front and -flow through a number of slag pots for separation. The shells which -remain adhering to the walls of the pots on pouring out the slag are -returned to the furnace. All the waste slag (containing about 0.5 per -cent. lead) is dumped down a ravine belonging to the territory of the -smeltery. - -The lead from the shaft furnace is liquated in a small reverberatory -furnace, of which the hearth consists of two inclined perforated -iron plates. The residue is returned to the shaft furnace, while the -liquated lead flows directly to the refining kettle, which is filled -in the course of four hours. Here it is steamed for about one hour and -is then cast into bars through a Steitz siphon, after skimming off the -oxide. The matte is crushed and roasted in a reverberatory furnace (60 -ft. long). - -The power plant comprises three Stirling boilers and two 250 h. p. -compound engines, of which one is for reserve; also one steam-driven -dynamo, coupled direct to the engine, furnishing the current for the -entire plant, for the electric locomotives, etc. - -The coke is obtained from Pennsylvania and costs about $4 a ton, while -the coal comes from near-by collieries and costs $1 per ton. - -In the well-equipped laboratory the lead in the ores and slags is -determined daily by Alexander’s (molybdate) method, while the silver -content of the lead (a little over 1 oz. per ton) is estimated only -once a month in an average sample. When the plant is in full operation -it gives employment to 150 men. Cases of lead-poisoning are said to -occur but rarely, and then only in a mild form. - - - - - LEAD SMELTING AT TARNOWITZ - - (September 23, 1905) - - -The account of the introduction of the Huntington-Heberlein process at -Tarnowitz, Prussia, published elsewhere in this issue, is of peculiar -interest inasmuch as it tells of the complete displacement by the new -process of one of the old processes of lead smelting which had become -classic in the art. The roast-reaction process of lead smelting, -especially as carried out in reverberatory furnaces, has been for a -long time decadent, even in Europe. Tarnowitz was one of the places -where it survived most vigorously. - -Outside of Europe, this process never found any generally extensive -application. It was tried in the Joplin district, and elsewhere in -Missouri, with Flintshire furnaces in the seventies. Later it was -employed with modified Flintshire and Tarnowitz furnaces at Desloge, -in the Flat River district of Missouri, where the plant is still in -operation, but on a reduced scale. - -The roast-reaction process of smelting, as practised at Tarnowitz, -was characterized by a comparatively large charge, slow roasting and -low temperature, differing in these respects from the Carinthian and -Welsh processes. It was not aimed to extract the maximum proportion of -lead in the reverberatory furnace itself, the residue therefrom, which -inevitably is high in lead, being subsequently smelted in the blast -furnace. Ores too low in lead to be suitable for the reverberatory -smelting were sintered in ordinary furnaces and smelted in the blast -furnace together with the residue from the other process. In both of -these processes the loss of lead was comparatively high. One of the -most obvious advantages of the Huntington-Heberlein process is its -ability to reduce the loss of lead. The result in that respect at -Tarnowitz is clearly stated by Mr. Biernbaum, whose paper will surely -attract a good deal of attention.[8] - - - - - LEAD SMELTING IN REVERBERATORY FURNACES AT DESLOGE, MO. - - BY WALTER RENTON INGALLS - - (December 16, 1905) - - -The roast-reaction method of lead smelting in reverberatory furnaces -never found any general employment in the United States, although -in connection with the rude air-furnaces it was early introduced in -Missouri. The more elaborate Flintshire furnaces were tried at Granby, -in the Joplin district, but they were displaced there by Scotch -hearths. The most extensive installation of furnaces of the Flintshire -type was made at Desloge, in the Flat River district of southeastern -Missouri. This continued in full operation until 1903, when the major -portion of the plant was closed, it being found more economical to ship -the ore elsewhere for smelting. However, two furnaces have been kept -in use to work up surplus ore. As a matter of historic interest, it is -worth while to record the technical results at Desloge, which have not -previously been described in metallurgical literature. - -The Desloge plant, which was situated close to the dressing works -connected with the mine, and was designed for the smelting of its -concentrate, comprised five furnaces. The furnaces were of various -constructions. The oldest of them was of the Flintshire type, and -had a hearth 10 ft. wide and 14 ft. long. The other furnaces were a -combination of the Flintshire and Tarnowitz types. They were built -originally like the newer furnaces at Tarnowitz, Upper Silesia, with a -rather large rectangular hearth and a lead sump placed at one side of -the hearth near the throat end; but good results were not obtained from -that construction, wherefore the furnaces were rearranged with the sump -at one side, but in the middle of the furnace, as in the Flintshire -form. The rectangular shape of the Tarnowitz hearth was, however, -retained. Furnaces thus modified had hearths 11 ft. wide and 16 ft. -long, except one which had a hearth 13 ft. wide. - -The same quantity of ore was put through each of these furnaces, the -increase in hearth area being practically of no useful effect, because -of inability to attain the requisite temperature in all parts of the -larger hearths with the method of heating employed. The men objected -especially to a furnace with hearth 13 ft. wide, which it was found -difficult to keep in proper condition, and also difficult to handle -efficiently. Even the width of 11 ft. was considered too great, and -preference was expressed for a 10 ft. width. In this connection, it may -be noted that the old furnaces at Tarnowitz were 11 ft. 9 in. long and -10 ft. 10 in. wide, while the new furnaces were 16 ft. long and 8 ft. -10 in. wide (Hofman, “Metallurgy of Lead,” fifth edition, p. 112). All -of these dimensions were exceeded at Desloge. - -The Flintshire furnaces at Desloge had three working doors per side; -the others had four, but only three per side were used, the doors -nearest the throat end being kept closed because of insufficient -temperature in that part of the furnace. The furnace with hearth 11 -× 14 ft. had a grate area of 6.5 × 3 ft. = 19.5 sq. ft.; the 11 × 16 -furnaces had grates 8 × 3 = 24 ft. sq. The ratios of grate to hearth -area were therefore approximately 1:8 and 1:7.3, respectively. (Compare -with ratio of 1:10 at Tarnowitz, and 1:6⅔ at Stiperstones.) The ash -pits were open from behind in the customary English fashion. The grate -bars were cast iron, 36 in. long. The bars were 1 in. thick at the top, -with ⅝ in. spaces between them. The open spaces were 32 in. long, -including the rib in the middle. The bars were 4 in. deep at the middle -and 2 in. at the ends. The distance from the surface of the grate bars -to the fire-door varied in the different furnaces. Some of those with -hearths 11 × 16 ft. and grates 8 × 3 ft. had the bars 6 in. below the -fire-door; in others the bars were almost on a level with the fire-door. - -The furnaces were run with a comparatively thin bed of coal on the -grate, and combustion was very imperfect, the percentage of unburned -carbon in the ash being commonly high. This was unavoidable with the -method of firing employed and the inferior character of the coal -(southern Illinois). The excessive consumption of coal was due largely, -however, to the practice of raking out the entire bed of coal at the -beginning of the operation of “firing down” (beginning the reaction -period), when a fresh fire was built with cordwood and large lumps of -coal. - -Each furnace had two flues at the throat, 16 × 18 in. in size, each -flue being provided with a separate damper. Each furnace had an -iron chimney approximately 55 ft. high, of which 13 ft. was a brick -pedestal (64 × 64 in.) and the remaining 42 ft. sheet steel, guyed. The -chimneys were 42 in. in diameter. The distance from the outside end -of the furnace to the chimney was approximately 6 ft., and there was -consequently but little opportunity for flue dust to collect in the -flue. About once a month, however, the chimney was opened at the base -and about two wheelbarrows (say 600 lb.) of flue dust, assaying about -50 per cent. lead, was recovered per furnace. - -The furnace house was a frame building 45 ft. wide, with boarded sides -and a corrugated-iron pitch roof, supported by steel trusses. The -furnaces were set in this house, side by side, their longitudinal axes -being at right angles to the longitudinal axis of the building. The -distance from the outside of the fire-box end of the furnace to the -side of the building was 10 ft. The coal was unloaded from a railway -track alongside of the building and was wheeled to the furnace in -barrows. Some of the furnaces were placed 18 ft. apart; others 22 ft. -apart. The men much preferred the greater distance, which made their -work easier, an important consideration in this method of smelting. - -The hight from the floor to the working door of the furnace was -approximately 36 in. The working doors were formed with cast-iron -frames, making openings 7 × 11 in. on the inside and 15 × 28 in. on -the outside. On the side of the furnace opposite the middle working -door was placed a cast-iron hemispherical pot, set partially below the -floor-line. This pot was 16 in. deep and 24 in. in diameter; the metal -was ¼ in. thick. The distance from the top of the pot to the line of -the working door was 31 in.; from the top of the pot to the bottom of -the tap-door was 7 in. The tap-door was 4 in. wide and 9 in. high, -opening through a cast-iron plate 1½ in. thick. Below the tap-door -and on a line with the upper rim of the pot was a tap-hole 3½ in. in -diameter. The frames of the working doors had lugs in front, against -which the buckstaves bore, to hold the frames in position. All other -parts of the sides of the furnace, including the fire-box, were cased -with ⅝ in. cast-iron plates, which were obviously too light, being -badly cracked. - -The cost of a furnace when built in 1893 was approximately $1400, -not including the chimney; but with the increased cost of material -the present expense would probably be about $2000. Notwithstanding -the light construction of the furnaces, repairs were never a large -item. Once a month a furnace was idle about 24 hours while the throat -was being cleaned out, and every two months some repairing, such as -relining the fire-boxes, etc., was required. If repairs had to be made -on the inside of the furnace, two days would be lost while it was -cooling sufficiently for the men to enter. In refiring a furnace, from -8 to 12 hours was required to raise it to the proper temperature. Out -of the 365 days of the year, a furnace would lose from 20 to 25 days, -for cleaning the throat and making repairs to the fire-box, arch, etc. - -When a furnace was run with two shifts the schedule of operation was as -follows: - - Drop charge 4 a.m. - Begin work 7 a.m. - Begin firing down 11 a.m. - Begin first tapping 1 p.m. - Rake out slag 2.30 p.m. - Begin second tapping 3 p.m. - Drop charge 4 p.m. - Begin working 5.30 p.m. - Begin firing down 11 p.m. - Begin first tapping 1 a.m. - Rake out slag 2.30 a.m. - Begin second tapping 3 p.m. - -With three shifts on a furnace, the schedule was as follows: - - Drop charge 7 a.m. - Begin firing down 12 a.m. - Begin tapping 1 p.m. - Rake out slag 2 p.m. - Begin tapping 2.30 p.m. - Drop charge 3 p.m. - Begin firing down 8 p.m. - Begin tapping 9 p.m. - Rake out slag 10 p.m. - Begin tapping 10.30 p.m. - Drop charge 11.00 p.m. - Begin firing down 4 a.m. - Begin tapping 5 a.m. - Rake out slag 6 a.m. - Begin tapping 6.30 a.m. - -The hearths were composed of about 8 in. of gray slag beaten down -solidly on a basin of brick, which rested on a filling of clay, rammed -solid. The hearth was patched if necessary after the drawing of each -charge. - -The system of smelting was analogous to that which was practiced -in Wales rather than to the Silesian, the charges being worked off -quickly, and with the aim of making a high extraction of lead directly -and a gray slag of comparatively low content in lead. The average -furnace charge was 3500 lb. At the beginning of the reaction period -about 85 to 100 lb. of crushed fluorspar was thrown into the furnace -and mixed well with the charge. The furnace doors were then closed -tightly and the temperature raised, the grate having previously been -cleaned. At the first tapping about 1200 lb. of lead would be obtained. -A small quantity of chips and bark was thrown into the lead in the -kettle, which was then poled for a few minutes, skimmed, and ladled -into molds, the pigs weighing 80 lb. The skimmings and dross were -put back into the furnace. The pig lead was sold as “ordinary soft -Missouri.” The gray slag was raked out of the furnace, at the end of -the operation, into a barrow, by which it was wheeled to a pile outside -of the building. Shipments of the slag were made to other smelters from -time to time, 95 per cent. of its lead content being paid for when its -assay was over 40 per cent., and 90 per cent. when lower. - -Each furnace was manned by one smelter ($1.75) and one helper ($1.55) -per shift, when two shifts per 24 hours were run. They had to get their -own coal, ore and flux, and wheel away their gray slag and ashes. In -winter, when three shifts were run, the men were paid only $1.65 and -$1.50 respectively. There was a foreman on the day shift, but none at -night. The total coal consumption was ordinarily about 0.8 to 0.9 per -ton of ore. Run-of-mine coal was used, which cost about $2 per ton -delivered. The coal was of inferior quality, and it was wastefully -burned, as previously referred to, wherefore the consumption was high -in comparison with the average at Tarnowitz, where it used to be about -0.5 per ton of ore. - -The chief features of the practice at Desloge are compared with those -at Tarnowitz, Silesia and Holywell (Flintshire), and Stiperstones -(Shropshire), Wales, in the following table, the data for Silesia and -Wales being taken from Hofman’s “Metallurgy of Lead,” fifth edition, -pp. 112, 113. - - ──────────────────────┬─────────┬────────┬─────────┬─────────┬──────── - DETAIL │HOLYWELL │ STIPER-│TARNOWITZ│TARNOWITZ│ DESLOGE - │ │ STONES │ │ │ - ──────────────────────┼─────────┼────────┼─────────┼─────────┼──────── - Hearth length, ft. │ 12.00 │ 9.75 │ 11.75 │ 16.00 │ 16.00 - Hearth width, ft. │ 9.50 │ 9.50 │ 10.83 │ 8.83 │ 11.00 - Grate length, ft. │ 4.50 │ 4.50 │ 8.00 │ 8.00 │ 8.00 - Grate width, ft. │ 2.50 │ 2.50 │ 1.67 │ 1.67 │ 3.00 - Grate area: hearth │ │ │ │ │ - area │ 1:8 │ 1:6⅔ │ 1:10 │ 1:10 │ 1:7⅓ - Charges per 24 hr., │ 3 │ 3 │ 2 │ 2 │ 3 - Ore smelted per │ │ │ │ │ - 24 hr., lb. │ 7,050 │ 7,050 │ 8,800 │ 16,500 │ 10,500 - Assay of ore, % Pb │ 75-80 │ 77.5 │ 70-74 │ 70-74 │ 70 - Gray slag, % of charge│ 12 │ │ 15 │ 30 │ 27 - Gray slag, % Pb │ 55 │ │ 38.8 │ 56 │ 38 - Men per 24 hr. │ 6 │ 4 │ 4 │ 6 │ 6 - Coal used per ton ore │0.57-0.76│ 0.56 │ 0.46 │ 0.50 │ 0.90 - ──────────────────────┴─────────┴────────┴─────────┴─────────┴──────── - -The regular furnace charge at Desloge was 3500 lb. The working of three -charges per 24 hours gave a daily capacity of 10,500 lb. per furnace. -These figures refer to the wet weight of the concentrate, which was -smelted just as delivered from the mill. Its size was 9 mm. and finer. -Assuming its average moisture content to be 5 per cent., the daily -capacity per furnace was about 10,000 lb. (5 tons) of dry ore. - -The metallurgical result is indicated by the figures for two months -of operation in 1900. The quantity of ore smelted was 1012 tons, -equivalent to approximately 962 tons dry weight. The pig lead produced -was 523.3 tons, or 54.4 per cent. of the weight of the ore. The gray -slag produced was 262.25 tons, or about 27 per cent. of the weight of -the ore. The assay of the ore was approximately 70 per cent. lead, -giving a content of 673.4 tons in the ore smelted. The gray slag -assayed approximately 38 per cent. lead, giving a content of 99.66 -tons. Assuming that 90 per cent. of the lead in the gray slag be -recoverable in the subsequent smelting in the blast furnace, or 89.7 -tons, the total extraction of lead in the process was 523.3 + 89.7 ÷ -673.4 = 91 per cent. The metallurgical efficiency of the process was, -therefore, reasonably high, especially in view of the absence of dust -chambers. - - * * * * * - -The cost of smelting with five furnaces in operation, each treating -three charges per day, was approximately as follows: - - 1 foreman at $3 $3.00 - 5 furnace crews at $9.90 49.50 - Unloading 21 tons of coal at 6c. 1.26 - Loading 14 tons lead at 15c. 2.10 - “ 7 tons gray slag at 15c. 1.05 - —————— - Total labor $56.91 - - 21 tons coal at $2 $42.00 - Flux and supplies 13.00 - Blacksmithing and repairs 10.00 - —————— - Total $121.91 - -On the basis of 6.25 tons of wet ore, this would be $4.65 per ton. The -actual cost in seven consecutive months of 1900 was as follows: Labor, -$1.98 per ton; coal, $1.86; flux and supplies, $0.51; blacksmithing and -repairs, $0.39; miscellaneous, $0,017; total, $4.757. If the cost of -smelting the gray slag be reckoned at $8 per ton, and the proportion -of gray slag be reckoned at 0.25 ton per ton of galena concentrate, -the total cost of treatment of the latter comes to about $6.75 per ton -of wet charge, or about $7 per ton of dry charge. This cost could be -materially reduced in a larger and more perfectly designed plant. - -The practice at Desloge did not compare unfavorably, either in respect -to metal extracted or in smelting cost, with the roast-reduction method -of smelting or the Scotch hearth method, as carried out in the plants -of similar capacity and approximately the same date of construction, -smelting the same class of ore, but the larger and more recent plants -in the vicinity of St. Louis could offer sufficiently better terms to -make it advisable to close down the Desloge plant and ship the ore to -them. One of the drawbacks of the reverberatory method of smelting -was the necessity of shipping away the gray slag, the quantity of -that product made in a small plant being insufficient to warrant the -operation of an independent shaft furnace. - - - - - PART III - - SINTERING AND BRIQUETTING - - - - - THE DESULPHURIZATION OF SLIMES BY HEAP ROASTING AT BROKEN HILL[9] - - BY E. J. HORWOOD - - (August 22, 1903) - - -It is well known that, owing to the intimate mixture of the -constituents of the Broken Hill sulphide ores, a great deal of crushing -and grinding is required to detach the particles of galena from the -zinc blende and the gangue; and it will be understood, therefore, that -a considerable amount of the material is converted into a slime which -consists of minute but well-defined particles of all the constituents -of the ore, the relative proportions of which depend on the dual -characteristics of hardness and abundance of the various constituents. -An analysis of the slime shows the contents to be as follows; - - Galena (PbS) 24.00 - Blende (ZnS) 29.00 - Pyrite (FeS₂) 3.38 - Ferric oxide (Fe₂O₃) 4.17 - Ferrous oxide (FeO) contained in garnets 1.03 - Oxide of manganese (MnO) contained in rhodonite and garnets 6.66 - Alumina (Al₂O₃) contained in kaolin and garnets 5.40 - Lime (CaO) contained in garnets, etc. 3.40 - Silica (SiO₂) 22.98 - Silver (Ag) .06 - —————— - 100.48 - -Galena, being the softest of these, is found in the slimes to a larger -extent than in the crude ore; it is also, for the same reason, in the -finest state of subdivision, as is well illustrated by the fact that -the last slime to settle in water is invariably much the richest in -lead, while the percentages of the harder constituents, zinc blende and -gangue, show a corresponding reduction in quantity, by reason of their -being generally in larger sized particles and consequently settling -earlier. - -The fairly complete liberation of each of the constituent minerals -of the ore that takes place in sliming tends, of course, to help -the production of a high-grade concentrate by the use of tables and -vanners, and undoubtedly a fair recovery of lead is quite possible, -even with existing machines, in the treatment of fine slimes; but, -owing to the great reduction in the capacity of the machines, which -takes place when it is attempted to carry the vanning of the finer -slimes too far, and the consequently greatly increased area of the -machines that would be necessary, the operation, sooner or later, -becomes unprofitable. - -The extent to which the vanner treatment of slimes should be carried -is, of course, less in the case of those mines owning smelters than -with those which have to depend on the sale of concentrates as their -sole source of profit. In the case of the Proprietary Company, -all slime produced in crushing is passed over the machines after -classification. A high recovery of lead in the form of concentrates -is, of course, neither expected nor obtained, for reasons already -explained; but the finest lead-bearing slimes are allowed to unite -with the tailings, which are collected from groups of machines, and -are then run into pointed boxes, where, with the aid of hydraulic -classification, the fine rich slimes are washed out and carried to -settling bins and tanks, where the water is stilled and allowed to -deposit its slime, and pass over a wide overflow as clear water. The -slime thus recovered amounts to over 1200 tons weekly, or about 11 per -cent., by weight, of the ore, and assays about 20 per cent. lead, 17 -per cent. zinc, and 18 oz. silver, and represents, in lead value, about -11 per cent. of the original lead contents of the crude ore and rather -more than that percentage in silver contents. These slimes are thus a -by-product of the mills, and their production is unavoidable; but as -they are not chargeable with the cost of milling, they are an asset of -considerable value, more especially so since it has been demonstrated -that they can be desulphurized sufficiently for smelting purposes by a -simple operation, and, at the same time, converted into such a physical -condition as renders the material well suited for smelting, owing to -its ability to resist pressure in the furnaces. - -The Broken Hill Proprietary Company has many thousands of tons of -these slimes which the smelters have hitherto been unable to cope with, -owing to the roasters being fully occupied with the more valuable -concentrates. Moreover, the desulphurization of slimes in Ropp -mechanical roasters is objectionable for various reasons, namely, owing -to the large amount of dust created with such fine material, resulting -injuriously to the men employed; also on account of the reduction in -the capacity of the roasters, and consequent increase in working cost, -owing to the lightness of the slime, especially when hot, as compared -with concentrates, and the necessity for limiting the thickness of -material on the bed of the roasters to a certain small maximum. -Further, the desulphurization of the slimes is no more complete with -the mechanical roasters than in the case of heap roasting, and the -combined cost of roasting and briquetting being quite three shillings -(or 75c.) per ton in excess of the cost of heap roasting, the -latter possesses many advantages. These heaps are being dealt with, -preparatory to roasting, by picking down the material in lumps of about -5 in. in thickness, while the fine dry smalls, unavoidably produced, -are worked up in a pug mill with water, and dealt with in the same way -as the wet slime produced from current work. - -The slime, as produced by the mills, is run from bins into railway -trucks in a semi-fluid condition, and shortly after being tipped -alongside one of the various sidings on the mine is in a fit condition -to be cut with shovels into rough bricks, which dry with fair rapidity, -and when required for roasting are easily reloaded into railway trucks. -As each man can cut about 20 tons of bricks per day, the cost is small. -Various other methods of lumping the slime were tried, including -trucking the semi-fluid material on movable trams, alongside which were -set laths, about 9 in. apart, which enabled long slabs to be formed -9 in. wide and 5 in. thick, which were, after drying, picked up in -suitable lumps and loaded in platform trucks, thence on railway trucks. -Owing to the inferior roasting that takes place with bricks having flat -sides, which are liable to come into close contact in roasting, and -to the rather high labor cost, this method was discontinued. Another -method was to allow the slime to dry partially after being emptied -from railway trucks, and to break it into lumps by means of picks; -but this method entailed the making of an increased amount of smalls, -besides taking up more siding room, owing to the extra time required -for drying, as compared with the method now in use. Ordinary bricking -machines could, of course, be used, but when the cost of handling the -slime before and after bricking is counted, the cost would be greater -than with the simple method now in use; the material being in too -fluid a condition for making into bricks until some time elapses for -drying, a double handling would be necessitated before sending it to -the bricking machine. If, however, the slime could be allowed time to -dry sufficiently in the trucks, bricking by machinery would probably be -preferable. Rather more than 10 per cent. of smalls is made in handling -the lumps in and out of the railway trucks, and this is, as already -noted, worked up with water in a pug mill at the sintering works, and -used partly for covering the heaps with slime to exclude an excessive -amount of air. The balance is thrown out and cut into bricks, as -already described. - -At the heaps the lumps are at present being thrown from one man to -another to reach their destination in the heap, but the sidings have -been laid out in duplicate with a view to enabling traveling cranes to -be used on the line next the heap, the lumps to be loaded primarily -into wooden skips fitting the trucks. It is probable, however, that -the lumps will require to be handled out of the skips into their place -in the heap, as the brittle nature of the material may be found to -render automatic tipping impracticable. A considerable saving in labor -would nevertheless accompany the use of cranes, which would likewise be -advantageous in loading the sintered material. - -In order to reduce the inconvenience arising from fumes, length is very -desirable in siding accommodation, so that heap building may be carried -on at a sufficient distance from the burning kilns. It is for the same -reason preferable to build in a large tonnage at one time, lighting -the heaps altogether. As the heaps burn about two weeks only, long -intervals intervene, during which the fumes are absent. - -In the experimental stages of slime roasting, fuel, chiefly wood, was -used in quantities up to 5 per cent., and was placed on the ground at -the bottom of the heap, where also a number of flues, loosely built -bricks, were placed for the circulation of air. The amount of fuel -used has, however, been gradually reduced, until the present practice -of placing no fuel whatever in the bottom was arrived at; but instead -less than 1 per cent. of wood is now burned in small enlargements of -the flues, under the outer portion of the pile, and placed about 12 -ft. apart at the centers. This is found to be sufficient to start the -roasting operation within 24 hours of lighting, after which no further -fuel is necessary. - -As regards the dimensions of the heaps, the width found most suitable -is 22 ft. at the base, the sides sloping up rather flatter than one to -one, with a flat section on top reaching about 7 ft. in hight. As there -is always about 6 in. of the outer crust imperfectly roasted, it is -advisable to make the length as great as possible, thus minimizing the -surface exposed. The company is building heaps up to 2000 ft. long. - -During roasting care is required to regulate the air supply, the object -being to avoid too fierce a roast, which tends to sinter and partially -fuse the material on the outer portions of the lumps, while inside -there is raw slime. By extending the roast over a longer period this is -avoided, and a more complete desulphurization is effected. Experiments -conducted by Mr. Bradford, the chief assayer, demonstrated that, at a -temperature of 400 deg. C., the sulphide slime is converted into basic -sulphate, while at a temperature of 800 deg. C. the material becomes -sintered owing to the decomposition of the basic sulphate and the -formation of fusible silicate of lead. - -In practice, the sulphur contents of the material, which originally -are about 14 per cent., become reduced to from 6.5 to 8.5 per cent., -half in the form of basic sulphate and half as sulphides; much of the -material sinters and becomes matted together in a fairly solid mass. -The heaps are built without chimneys of any kind; a strip about 5 -ft. wide along the crest of the pile is left uncovered by plastered -slime, and this, together with the open way in which the lumps are -built in, allows a natural draft to be set up, which can be regulated -by partly closing the open ends of the flues at the base of the pile. -Masonry kilns were used in the earlier stages with good results, which, -however, were not so much better than those obtained by the heap method -as to justify the expense of building, taking into consideration, too, -the extra cost of handling the roasted material in the necessarily more -confined space. - -Much interest has been taken in the chemical reactions which take -place in the operation of desulphurization of these slimes, it being -contended, on the one hand, that the unexpectedly rapid roast which -takes place may be due to the sulphide being in a very fine state of -subdivision, and more or less porous, thus allowing the air ready -access to the sulphur, producing sulphurous acid gas (SO₂). On the -other hand, others, of whom Mr. Carmichael is the chief exponent, claim -that several reactions take place during the operation, connected -with the rhodonite and lime compounds present in the slimes, which he -describes as follows: - -“The temperature of the kilns having reached a dull red heat, the -rhodonite (silicate of manganese) is converted into manganous oxide -and silica; at a rather higher temperature the calcium compounds are -also split up, with formation of calcium sulphide, the sulphur being -provided by the slimes. The air permeating the mass oxidizes the -manganese oxide and calcium sulphide into manganese tetroxide and -calcium sulphate respectively, as shown as follows; - - 3MnO + O = Mn₃O₄ - CaS + 4O = CaSO₄, - -and, as such, are carriers of a form of concentrated oxygen to the -sulphide slimes, with a corresponding reduction to manganous oxide and -calcium sulphide, as shown by the following equation, in the case of -lead: - - PbS + 4Mn₃O₄ = PbSO₄ + 12MnO - PbS + CaSO₄ = PbSO₄ + CaS. - -The oxidation of the manganous oxide and calcium sulphide is repeated, -and these alternate reactions recur until the desulphurization ceases, -or the kiln cools down to a temperature below which oxidation cannot -occur. These reactions, being heat-producing, provide part of the heat -necessary for desulphurization, which is brought about by certain -concurrent reactions between metallic sulphates and sulphide. - -“The first that probably occurs is that in which two equivalents of the -metallic sulphide react on one of the metallic sulphate with reduction -to the metal, metallic sulphide, and sulphurous acid, as shown by the -following equation in the case of lead: - - 2PbS + PbSO₄ = 2Pb + PbS + 2SO₂. - -“The metal so formed, in the presence of air, is oxidized, and in this -state reacts on a further portion of the metallic sulphide produced, -with an increased formation of metal and evolution of sulphurous acid, -according to the following equation, in the case of lead: - - 2PbO + PbS = Pb + SO₂. - -“The metal so produced in this reaction is wholly reoxidized by the -oxygen of the air current, and being free to react on still further -portions of the metallic sulphide, repeats the reaction, and becomes -an important factor in the desulphurizing of the undecomposed portion -of the material. As the desulphurization proceeds, and the sulphate of -metal accumulates, reactions are set up between the metallic sulphide -and different multiple proportions of the metallic sulphate, with the -formation of metal, metallic oxide, and evolution of sulphurous acid, -as follows: - -“With two equivalents of metallic sulphate to one equivalent of -metallic sulphide, in the case of lead, according to the following -equation: - - PbS + 2PbSO₄ = 2PbO + Pb + 3SO₂. - -“With three equivalents of metallic sulphate to one of metallic -sulphide, in the case of lead, according to the following equation: - - PbS + 3PbSO₄ = 4PbO + 4SO₂.” - -The volatility of sulphide of lead—especially in the presence of an -inert gas such as sulphurous acid—being greater than that of the -sulphate, oxide, or the metal itself, it might be thought that the -conditions are conducive to a serious loss of lead. This, however, is -reduced to a minimum, owing to the easily volatilized sulphide being -trapped, as non-volatile sulphate, by small portions of sulphuric -anhydride (SO₃), which is formed by a catalytic reaction set up -between the hot ore, sulphurous acid, and the air passing through -the mass. Owing to the non-volatility of the silver compounds in the -slimes, the loss of this metal has been found to be inappreciable. The -zinc contents of the slime are reduced appreciably, thus rendering the -material more suitable for smelting. After desulphurization ceases, -a few days are allowed for cooling off. On the breaking up of the -mass for despatch to the smelters, as much of the lower portion of -the walls is left intact as possible, so that it can be utilized for -the next roast, thus avoiding the re-building of the whole of the -walls.[10] - - - - - THE PREPARATION OF FINE MATERIAL FOR SMELTING - - BY T. J. GREENWAY - - (January 12, 1905) - - -In the course of smelting, at the works of the company known as the -Broken Hill Proprietary Block 14, material which consisted chiefly of -silver-lead concentrate and slime, resulting from the concentration -of the Broken Hill complex sulphide ore, I had to contend with all -the troubles which attend the treatment of large quantities of finely -divided material in blast furnaces. With the view of avoiding these -troubles, I experimented with various briquetting processes; and, -after a number of more or less unsatisfactory experiences, I adopted a -procedure similar to that followed in manufacturing ordinary bricks by -what is known as the semi-dry brick-pressing process. This method of -briquetting not only converts the finely divided material cheaply and -effectively into hard semi-fused lumps, which are especially suitable -for the heavy furnace burdens required by modern smelting practice, but -also eliminates sulphur, arsenic, etc., to a great extent; therefore, -it is capable of wide application in dealing with concentrate, slime, -and other finely divided material containing lead, copper and the -precious metals. - -This briquetting process comprises the following series of operations: - -1. Mixing the finely divided material with water and newly slaked lime. - -2. Pressing the mixture into blocks of the size and shape of ordinary -bricks. - -3. Stacking the briquettes in suitably covered kilns. - -4. Burning the briquettes, so as to harden them, without melting, at -the same time eliminating sulphur, arsenic, etc. - -1. The material is dumped into a mixing plant, together with such -proportions of screened slaked lime (usually from three to five per -cent.) and water as shall produce a powdery mixture which will, on -being squeezed in the hand, cohere into dry lumps. In preparing the -mixture, it is well to mix sandy material with suitable proportions -of fine, such as slime, in order that the finer material may act as a -binding agent. - -The mixer used by me consists of an iron trough, about 8 ft. long, -traversed by a pair of revolving shafts, carrying a series of knives -arranged screw-fashion; and so placed that the knives on one shaft -travel through the spaces between the knives on the other shaft. -The various materials are dumped into one end of the mixing trough, -from barrows or trucks, and are delivered continuously at the other -end of the trough, into an elevator which conveys the mixture to the -brick-pressing plant. - -2. The plant employed was the semi-dry brick-press. This machine -receives the mixture from the elevators, and delivers it in the form -of briquettes, which can at once be stacked in the kilns. It was found -that such material as concentrate and slime has comparatively little -mobility in the dies during the pressing operation; this necessitates -the use of a device which provides for the accurate filling of the -dies. It was also found that the materials treated by smelters vary -in compressibility, and this renders necessary the adoption of a -brick-pressing plant having plungers which are forced into the dies by -means of adjustable springs, brick-presses having plungers actuated by -rigid mechanism being extremely liable to jam and break. - -3. Briquettes made from such material as concentrate and slime vary -in fusibility; they are also combustible, and while being burned they -produce large quantities of smoke containing sulphurous acid and other -objectionable fumes. It is therefore necessary that such briquettes be -burned in kilns provided with arrangements for accurately controlling -the burning operations, and for conveniently disposing of the smoke. -Suitable kilns, which will contain from 30 to 50 tons of briquettes -per setting, are employed for this purpose. Regenerative kilns of the -Hoffman type might be used for dealing with some classes of material, -but, for general purposes, the kilns as designed here will be found -more convenient. - -The briquettes are stacked according to the character of the material -and the object to be obtained. The various methods of stacking, and the -reasons for adopting them, can be readily learned by studying ordinary -brick-burning operations in any large brick-yard. After the stacking -is complete the kiln-fronts are built up with burnt briquettes produced -in conducting previous operations, and all the joints are well luted. - -4. In burning briquettes made from pyrite or other self-burning -material, it is simply necessary to maintain a fire in the kiln -fireplaces for a period of from 10 to 20 hours. When it is judged that -this firing has been continued long enough, the fire-bars are drawn -and the fronts are luted with burnt briquettes in the same manner as -the kiln-fronts. Holes about two inches square are then made in these -lutings, through which the air required for the further burning of the -briquettes is allowed to enter the kilns under proper control. After -the fireplaces are thus closed the progress of the burning, which -continues for periods of from three to six days, is watched through -small inspection holes made in the kiln-fronts; and when it is seen -that the burning is complete the fronts are partially torn away, -in order to accelerate the cooling of the burnt briquettes, which -are broken down and conveyed to the smelters as soon as they can be -conveniently handled. - -When briquettes made from pyrite concentrate, or of other free-burning -material, are thus treated, they are not only sintered but they are -also more or less effectively roasted, and it may be taken for granted -that any ore which can be effectively roasted in the lump form in kilns -or stalls will form briquettes that will both sinter and roast well; -indeed, one may say more than this, for briquettes which will sinter -and roast well can be made from many classes of ore that cannot be -effectively treated by ordinary kiln-and stall-roasting operations; -and, moreover, good-burning briquettes may be made from mixtures of -free-burning and poor-burning material. Briquettes containing large -proportions of pyrite or other free-burning material will, unless the -air-supply is properly controlled, often heat up to such an extent as -to fuse into solid masses, much in the same manner as matte of pyritic -ore will melt when it is unskilfully handled in roasting. In dealing -with material which will not burn freely, such as roasted concentrate, -the briquetting is conducted with the intention of sintering the -material; and in this case the firing of the kilns is continued for -periods of from three to four days, the procedure being similar in -every way to that followed in burning ordinary bricks. - -When conducting my earlier briquetting operations I made the -briquettes by simply pugging the finely divided material, following -a practice similar to that adopted in producing “slop-made” bricks -by hand. This method of making the briquettes was attended with a -number of obvious disadvantages, and was abandoned as soon as the -semi-dry brick-pressing plant became available. The extent to which -this process, or modifications of it, may be applied is shown by the -fact that, following upon information given by me, the Broken Hill -Proprietary Company adopted a similar method of sintering and roasting -slime, consisting of about 20 per cent. galena, 20 per cent. blende, -and 60 per cent. silicious gangue. The procedure followed in this -case consisted of simply pugging the slime, and running the pug upon -a floor to dry; afterward cutting the dried material into lumps by -means of suitable cutting tools, and then piling the lumps over firing -foundations, following a practice similar to that pursued in conducting -ordinary heap-roasting. This company is now treating from 500 to 1000 -tons of slime weekly in this manner. It is, however, certain that -better results would attend the treatment of this material by making -this slime into briquettes and burning them in kilns. - -The cost of briquetting and burning material in the manner first -described, with labor at 25c. per hour, and wood or coal at $4 per ton, -amounts to from $1 to $1.50 per ton of material. - - - - - THE BRIQUETTING OF MINERALS - - BY ROBERT SCHORR - - (November 22, 1902) - - -The value of briquetting in connection with metallurgical processes and -the manufacture of artificial stone is well understood and appreciated. -In smelting plants there is always more or less flue dust, fine ores, -and sometimes fine concentrates to be treated, but the charging -of such fine material directly into a furnace would cause trouble -and irregularities, and would lessen its capacity also. As mineral -briquetting cannot be effected without considerable wear upon the -machinery and without quite appreciable expense in binder, labor, and -handling, many smelters try to avoid it. - -The financial question, however, is not as serious as it may at first -appear, and taking the large output of modern briquetting machines in -consideration, the cost for repairs amounts only to a few cents per ton -of briquetted material. The total cost depends in the first place on -the cost of labor, power and the binder, and in most American smelters -it varies between $0.65 and $1.25 per ton of briquettes. - -Ordinary brick presses, with clay as a binder, were used in Europe as -well as in this country, but they are too slow and expensive for large -propositions and the presence of clay is usually undesirable. - -The English Yeadon (fuel) press has also been used for some years at -the Carlton Iron Company’s Works at Ferryhill in England, and at the -Ore and Fuel Company’s plant at Coatbridge in the same country; also by -some Continental firms. Dupuis & Sons, Paris, furnished a few presses -which are mostly used for manganese and iron ores and pyrites. In -some localities coke dust is added. The making of clay briquettes or -mud-cakes is the crudest form of briquetting; but while heat has to -be expended to evaporate the 40 to 50 per cent. of moisture in them, -and while considerable flue dust is made, this method is better than -feeding fine ore or flue dust directly into the furnace. - -The only other method of avoiding briquetting is by fusing ore fines in -slagging reverberatory furnaces and by adding flue dust in the slagging -pit, thus incorporating it with the slagging ore. This is practised -sometimes in silver-lead smelters, but in connection with copper or -iron smelters it is not practicable. - -In briquetting minerals a thorough mixing and kneading is of the first -importance. If this is done properly a comparatively low pressure will -suffice to create a good and solid briquette, which after six to eight -hours of air-drying, or after a speedier elimination of the surplus of -moisture in hot-air chambers, will be ready for the furnace charge. A -good briquette should permit transportation without excessive breakage -or dust a few hours after being made, and it should retain its shape in -the furnace until completely fused, so as to create as little flue dust -as possible. The briquette should be dense, otherwise it will crumble -under the influence of bad weather. - -The two presses on the American machinery market are the type built by -the Chisholm, Boyd & White Company, of Chicago, and the briquetting -machine manufactured by the H. S. Mould Company, of Pittsburg. Both are -extensively used, and in many metallurgical plants it will pay well to -adopt them. - -From 4 to 6 per cent. of milk of lime is generally used as binder, -and this has a desirable fluxing influence also. A complete outfit -comprises, besides the press, a mixer for slacking the lime, and a -feed-pump which discharges the liquid in proportion into the main mixer -wherein the ore fines, flue dust, or concentrates are shoveled. - -The Chisholm, Boyd & White Company’s press makes 80 briquettes per -minute, which, with a new disk, are of 4 in. diameter and 2½ in. hight, -thus giving about 872 cu. ft. of briquette volume per 10 hours, or 50 -to 80 tons, depending on the weight of the material. With the wear of -the disk the hight of the briquettes is reduced and consequently the -capacity of the machine also. The disk weighs about 1600 lb., and as -most large smelters have their own foundries it can be replaced with -little expense. About 30 effective horse-power is usually provided for -driving the apparatus. The machine is too well known to metallurgists -and engineers to require further comment or description. - -The H. S. Mould Company has also succeeded in making its machine a -thorough practical success. This machine is a plunger-type press. The -largest press built employs six plungers, and at 25 revolutions it -makes 150 briquettes of 3 in. diameter and 3 in. hight, or 1080 cu. ft. -per 10 hours. Its rated capacity is 100 tons per 10 hours. - -In using a plunger-type press the material should not contain more -than 7 per cent. mechanical moisture. If wet concentrates have to -be briquetted it is necessary to add dry ore fines or flue dust to -arrive at a proper consistency. The briquettes are very solid and only -air-drying for a few hours is necessary. - -The cylindrical shape of briquettes is very good, as it insures -a proper air circulation in the furnace and consequently a rapid -oxidation and fusion. - -The wear of the Mould Company’s press is mostly confined to the chilled -iron bushings and to the pistons. Auxiliary machinery consists of -the slacker, the feeder and the main mixer. The press is of a very -substantial design, and it is claimed that the cost of repairs does not -amount to more than 3c. per ton of briquettes. - -Wear and tear is unavoidable in a crude operation like briquetting; to -treat flue dust, ore fines, and fine concentrates successfully, it is -almost absolutely necessary to resort to it. - -Edison used a number of intermittent-acting presses at his magnetic -iron-separation works in New Jersey, but this plant shut down some time -ago. - - - - - A BRICKING PLANT FOR FLUE DUST AND FINE ORES - - BY JAMES C. BENNETT - - (September 15, 1904) - - -The plant, which is here described, for bricking fine ores and flue -dust, was designed and the plans produced in the engineering department -of the Selby smelter. The machinery contained in the plant consists of -a Boyd four-mold brick press, a 7 ft. wet pan or Chile mill, a 50 h.p. -induction motor, and a conveyor-elevator, together with the necessary -pulleys and shafting. - -The press, Chile mill, and motor need no special mention, as they all -are from standard patterns and bought, without alterations, from the -respective builders. The Chile mill was purchased from the builders -of the brick press. The conveyor-elevator was built on the premises -and consists of a 14 in. eight-ply rubber belt, with buckets of sheet -steel placed at intervals of 6 in., running over flanged pulleys. The -buckets, or more properly speaking the flights, are made from No. -12 steel plate, flanged to produce the back and ends, with the ends -secured to the flanged bottom by one rivet in each. The plant has been -in operation for sixteen months and there have been few or no repairs -to the elevator, except to renew the belt, which is attacked by the -acid contained in the charges. This first belt was in continuous use -for nine months. As originally designed, the capacity was 100 tons per -day of 12 hours, but this was found to require a speed so high that -the workmen were unable to handle the output of the press. The speed -was, consequently, reduced about 25 per cent., which brings the output -down to about 75 tons per day. This output, as expressed in weight, -naturally varies somewhat owing to the variation in the weight of the -material handled. - -It is probable that the capacity could be increased to about 90 tons -by enlarging the bricks, which could be done, but would require a -considerable amount of alteration in the machine, as it is designed to -produce a standard sized building brick. By this method of increase, -however, the work of handling would not be materially increased, -because the number of bricks would be the same as with the present -output of 75 tons; there would be about 16 per cent. more to handle, -by weight. Working on the basis of 100 tons capacity, the bins were -designed to afford storage room for about three days’ run, or a little -over 300 tons. The bins are made entirely of steel, in order that -the hot material may be dumped into them directly from the roasting -furnaces, thus saving one handling. In order that there may be room -for several kinds of material, the bins are divided into seven -compartments, three on one side and four on the other. The lower part -is of ⅜ in. steel plate, and the upper, about one-half the hight, of -5/16 in. plate. - -It may be well to call attention to the method of handling the -material, preparatory to its delivery to the brick press. The bins are -constructed, as will be seen by the drawing, with their floor set 2.5 -ft. above the working floor, which enables the workmen to reach the -material with a minimum effort. The floor of the bins project 2.5 ft. -in front of the face, thus forming a platform on which the shoveling -may be done without the necessity of bending over. In this projecting -platform are cut rectangular holes 12 × 18 in., which are placed -midway between the openings in the front of the bins and furnished -with screens to stop any stray bolts or other coarse material that -might injure the press. This position of the holes through the platform -was adopted so that, in the event of the material running out beyond -the opening in the face, it would not fall directly upon the floor. -Two buckets are provided, with a capacity of 7 cu. ft. each, which is -the size of a single charge of the Chile mill. These buckets have a -hopper-shaped bottom fixed with a swinging gate which is operated by -the foot; thus the bucket can be run over the pan of the Chile mill and -the charge dumped directly into it. The buckets run on an overhead iron -track (1 in. by 3 in.) hung 7 ft. in the clear, above the floor. - -The method of making up the charge is as follows: The bucket is -run under the hole in the platform nearest to the compartment -containing the material of which the charge is partly composed, and -a predetermined number of shovelfuls is drawn out and put into the -bucket, which is then pushed on to the next compartment from which -material is wanted, where the operation is repeated. After charging -into the bucket the requisite amount of ore or flue dust, the bucket -is run to the back of the building, where the necessary amount of lime -(slaked) is added. By putting the lime in last, it is so surrounded by -the dust or ore that it has not the opportunity to stick to the sides -of the bucket in discharging, as it otherwise would. - -[Illustration: FIG. 1 (_a_).—Plant for Bricking Ores, Selby Smelter. -(Plan.)] - -The number of men required to operate the entire plant, exclusive -of those employed in bringing the material to the bins and emptying -the cars into them, is 12, placed as follows; One preparing the lime -for use, one removing the charge from the mill and supplying the -elevator-conveyor, which is accomplished by means of a specially -shaped, long-handled shovel; one keeping the supply spout of the press -clear (an attempt was made to do this mechanically, but was found to be -unsuccessful, owing to the extremely sticky nature of the material, and -so was discarded in favor of manual labor); one to control the press in -case of mishap and to keep the dies clean; one oiler; three receiving -the bricks from the press and taking the brick-loaded cars from the -press to the drying-house, and two placing the bricks on the shelves. - -[Illustration: FIG. 1 (_b_).—Plant for Bricking Ores, Selby Smelter. -(Elevation.)] - -The drying-house scarcely requires description; it is but a roofed -shed, without sides, fitted with stalls into which the bricks are set -on portable shelves, as close as working conditions will permit. The -means of drying, at the present time, is by the natural circulation -of air, but a mechanical system is in contemplation, by which the -air will be drawn into the building from the outside and forced to -find its way out through the bricks. The drying-house is adjacent to -the pressing plant, in fact forms the back of it, so that there is a -minimum distance to haul the product. The time required for drying the -bricks sufficiently for them to withstand the necessary handling is, -depending on the weather, from two to eight days, the usual time being -about three days. - - - - - PART IV - - SMELTING IN THE BLAST FURNACE - - - - - MODERN SILVER-LEAD SMELTING[11] - - BY ARTHUR S. DWIGHT - - (January 10, 1903) - - -The rectangular silver-lead blast furnace developed in the Rocky -Mountains has an area of 42 × 120 to 48 × 160 in. at the tuyeres; 54 -× 132 to 84 × 200 in. at the top; and hight from tuyere level to top -of charge of 15 to 21 ft. Such a furnace smelts 80 to 200 tons of -charge (ore and flux, but not slag and coke) per 24 hours. The slag -that has to be resmelted amounts to 20 to 60 per cent. of the charge. -Coke consumption is 12 to 16 per cent. of the charge. The blast -pressure ranges from 1.5 to 4 lb. per square inch, averaging close to -2 lb. Gases of hand-charged furnaces are taken off through an opening -below the charge-floor, the furnace being fed through a slot (about -20 in. wide, extending nearly the whole length of the furnace) in the -iron floor-plates; or through a hood (brick or sheet iron) above the -charge-floor level, with a down-take to the flues, charge-doors being -provided on each side of the hood, extending preferably the whole -length of the furnace and usually having a sill a few inches high which -compels the feeder to lift his shovel. - -When a silver-lead blast furnace is operating satisfactorily, the -following conditions should obtain; (1) A large proportion of the lead -in the charge should appear as direct bullion-product at the lead-well. -(2) The slag should be fluid and clean. (3) The matte should be low -in lead. (4) The furnace should be cool and quiet on top, making a -minimum quantity of lead-fume and flue-dust, and the charges should -descend uniformly over the whole area of the shaft. (5) The furnace -speed should be good. (6) The furnace should be free from serious -accretions and crusts; that is to say, the tuyeres should be reasonably -bright and open, and the level of the lead in the lead-well should -respond promptly to variations of pressure, caused by the blast and by -the hight of the column of molten slag and matte inside the furnace—an -indication that ample connection exists between the smelting column and -the crucible. Good reduction (using that term to express the degree in -which the furnace is manifesting its reducing action) is obtained when -the first three of the above conditions are satisfied. - -For any given furnace there are five prime factors, the resultant of -which determines the reduction, namely: (_a_) Chemical composition -of the furnace charges; (_b_) proportion and character of fuel; -(_c_) air-volume and pressure, to which might perhaps also be added -temperature of blast; for, although hot blast has not yet been -successfully applied in lead-smelting practice, I believe it is only -a question of time when it will be; (_d_) dimensions and proportions -of smelting furnace; (_e_) mechanical character and arrangement of the -smelting column. - -All but one of the above factors can be intelligently gaged. The -mechanical factor, however, can be expressed only in generalities and -indefinite terms. A wise selection of ores and proper preliminary -preparation, crushing the coarse and briquetting the fine, will do -much to regulate it, but all this care may be largely nullified by -careless feeding. The importance and possibilities of the mechanical -factor are generally overlooked and its symptoms are wrongly diagnosed. -For instance, the importance of slag-types has undoubtedly been -considerably exaggerated at the expense of the mechanical factor. -Slags seldom come down exactly as figured. We must know our ores and -apply certain empirical corrections to the iron, sulphur, etc., based -on previous experience with the ores; but these empirical corrections -may represent also an unformulated expression of the influence of the -mechanical factor on the reduction—a function, therefore, of the ruling -physical complexion of the ores, and the peculiarities of the feeding -habitually maintained in the works concerned. With a given ore-charge -large reciprocal variations may be produced in the composition of -slag and matte by merely changing the mechanical conditions of the -smelting column, and since the efficient utilization of both fuel and -blast must be controlled in the same way, the mechanical factor may be -considered, perhaps, the dominating agent of reduction. Inasmuch as -there is no way of gaging it, however, the only recourse is to seek a -correct adjustment and maintain it as a positive constant, after which -slag, fuel and blast may be with much greater certainty adjusted toward -efficiency of furnace work and metal-saving. - -_Behavior of Iron._—The output of lead is so dependent upon the -reactions of the iron in the charge that the chief attention may well -be fixed upon that metal as the key to the situation. The success of -the process depends largely upon reducing just the right amount of -iron to throw the lead out of the matte, the remainder of the iron -being reduced only to ferrous oxide and entering the slag. Too much -iron reduced will form a sow in the hearth. Iron is reduced from its -oxides principally by contact with solid incandescent carbon, and by -the action of hot carbon monoxide. Reduction by solid carbon is the -more wasteful, but there is in lead smelting an even more serious -objection to permitting the reduction to be accomplished by that means, -which leads to comparatively hot top and more or less volatilization of -lead. Reduction by carbon monoxide is the ideal condition for the lead -furnace. It means keeping the zone of incandescence low in the charge -column, leaving plenty of room above for the gases to yield up their -heat to, and exercise their reducing power on, the descending charge, -so that by the time they escape they will be well-nigh spent. Their -volume and temperature will be diminished, and the low velocity of -their exit will tend to minimize the loss of lead in fume and flue dust. - -The idea that high temperatures in lead blast furnaces should be -avoided is based on a misconception. Temperatures must exist which -are sufficiently high to volatilize all the lead in the charge, if -other conditions permit. A high temperature before the tuyeres means -fast smelting; and fast smelting, under proper conditions, means a -shortening of the time during which the lead is subject to scorifying -and volatilizing influences. A rapidly descending charge, constantly -replenished with cold ore from above, absorbs effectively the heat of -the gases and acts as a most efficient dust and fume collector. In -considering long flues, bag-houses, etc., it should be kept in mind -that the most effective dust collector ought to be the furnace itself. - -In the practice of twelve years ago and earlier, particularly when -using mixed coke and charcoal, reduction by carbon was probably the -rule; and the percentage of fuel required was very high. There is good -reason to think we have still much room for improvement along this line -in our average practice of today. - -_Volume of Blast._—It is customary to supply a battery of furnaces -from a large blast main, connected with a number of blowers. Inasmuch -as the air will take preferably the line of least resistance, if the -internal resistance of any one furnace be increased the volume of air -it will take will be diminished and the others will be favored unduly. -Only by keeping all the furnaces on approximately the same charge, with -the same hight of smelting column, can anything like uniformity of -operation and close regulation be secured. The rational plan would seem -to be to have a separate blower, of variable speed, directly connected -to each furnace, but this plan, which has had a number of trials, has -usually been abandoned in favor of the common blast main. Trials by -myself, extending over considerable periods, have been so uniformly -favorable, however, that I am forced to ascribe the failure of others -to some outside reason. - -The peculiar atmosphere required in the lead blast furnace depends -upon the correct proportion of two counteractive elements, carbon and -oxygen. If given too much air the furnace will show signs of deficient -reduction, commonly interpreted as calling for more fuel, which will -be sheer waste since its object is to burn up surplus air. There will -be an additional waste through the extra coal burned under the steam -boilers. The true remedy would be to cut down the quantity of air. -Burning up excessive coke is as hard work as smelting ore. Too much -fuel invariably slows up a furnace; it also drives the fire upward and -gives predominance to reduction by solid carbon. The maintenance of a -minimum fuel percentage, with a correctly adjusted volume of air, will -tend to promote the conditions under which iron will be reduced by the -gases, rather than by solid carbon. - -_Pressure of Blast._—Pressure necessarily involves resistance; and -the blast-pressure, as registered by a simple mercury-gage on the -bustle-pipe, may be increased in two ways: (1) By increasing the volume -of air forced through the interstices in the charge. This is the -wrong way; but, unfortunately, it is only too common in our practice, -and therefore deserves to be mentioned, if only to be condemned. (2) -By leaving the volume of air unchanged, but increasing the friction -offered by the interstitial channels, either by making them smaller in -aggregate cross-section (which means a finer charge), or by making them -longer (which means a higher smelting column). A correctly graduated -internal resistance is, therefore, the only true basis for a high blast -furnace, which, when so produced, will bring about rapid smelting, a -low zone of incandescence, and a very vigorous action upon the ores by -the gases in their retarded ascent through the charge column. These -conditions promote the reduction of iron by CO. The adjustment of -internal resistance, which is thus clearly the main factor, can be -accomplished only by the correct feeding of the furnace. - -_Feeding the Charge._—It is self-evident that, the more thorough the -preliminary preparation of the charge before it reaches the zone of -fusion, the more rapidly can the actual smelting proceed. A piece of -raw ore that finds itself prematurely at the tuyeres, without having -been subjected to the usual preparatory processes of drying, heating, -reduction, etc., must remain there until it is gradually dissolved or -carried away mechanically in the slag. Any such occurrence must greatly -retard the process. It would seem, by the same reasoning, that an -intimate mixture of the ingredients of the charge should expedite the -smelting, and I advocate the intimate mixture of the charge ingredients -in all cases. - -The theory of feeding is simple, but not so the practice. If the -charge column were composed of pieces of uniform size, the ascending -gases would find the channel of least resistance close to the furnace -walls and would take it preferably to the center of the shaft. The -more restricted channel would necessitate a higher velocity, so that -not only would the center of the charge be deprived of the action of -the gases, but also the portion traversed would be overheated; many -particles of ore would be sintered to the walls or carried off as flue -dust; slag would form prematurely; fuel would be wasted; in short, -all the irregularities and losses which accompany over-fire would be -experienced. In practice the charge is never uniform, but is a mixture -of coarse and fine. By lodging the finer material close to the walls -and placing the coarser in the center, an adjustment may be made which -will cause the gases to ascend uniformly through the smelting column. -A furnace top smoking quietly and uniformly over its whole area is the -visible sign of a properly fed furnace. - -_Effect of Large Charges._—It has frequently been remarked that, -within certain limits, large charges give more favorable results -than small ones; and numerous attempts have been made to account -for this fact. My observations lead me to offer the following as a -rational explanation—at least in cases where ore and fuel are charged -in alternate layers. Large ore-charges mean correspondingly large -fuel-charges. The gases can pass readily through the coke; and hence -each fuel-zone tends to equalize the gas currents by giving them -another opportunity to distribute themselves over the whole furnace -area, while each layer of ore subsequently encountered will blanket the -gases, and compel them to force a passage under pressure, which is the -manner most favorable to effective chemical action. - -In mechanically fed furnaces the charges of ore and fuel are usually -dropped in simultaneously from a car and the separate layers thus -obliterated, and the distributing zones which are such a safeguard -against the consequences of bad feeding are lacking, hence more care -must be exercised to secure proper placing of the coarse and fine -material. This may throw some light on the failure of most of the early -attempts at mechanical feeding. - -_Mechanical Character of Charge._—Very fine charges blanket the gases -excessively and cause them to break through at a few points, forming -blow-holes, which seriously disturb the operation, cause loss of raw -ore in the slag, and are accompanied by all the evils of over-fire. A -charge containing a few massive pieces, the rest being fine, is a still -more unfavorable combination. A very coarse charge permits too ready an -exit to the gases, and in the end tends likewise to over-fire and poor -reduction. The remedy is to briquette the fine ore (though preferably -not all of it), and crush the coarse to such degree as to approach an -ideal result, which may be roughly described as a mixture in which -about one-third is composed of pieces of 5 to 2 in. in diameter, -one-third pieces of 2 to 0.5 in., and the remaining third from 0.5 in. -down. The coke is better for being somewhat broken up before charging, -and a reasonable amount of coke fines, such as usually accompanies -a good quality of coke, is not in the least detrimental. The common -practice of handling the coke by forks and throwing away the fines -is to be condemned as an unwarranted waste of good fuel. The slag on -the charge should be broken to pieces at most 6 in. in diameter. The -common practice of throwing in whole butts of slag-shells is bad. -There is no economy in using the slag hot; cold charges, not hot, -are what we want. A reasonable amount of moisture in the charge is -beneficial, providing it be in such form as to be readily dried out. It -is often advantageous to wet the ore mixtures while bedding them, or -to sprinkle the charges before feeding. The driving off of this water -must consume fuel, but not so much as if the smelting zone crept up. -Large doses of water applied directly to the furnace are unpardonable -under any circumstances, however, though they are sometimes indulged -in as a drastic measure to subdue excessive over-fire when other and -surer means are not recognized. One of the chief merits of moderate -sprinkling before charging is that it gives in many cases a more -favorable mechanical character, approximating a lumpy condition in too -fine a charge, and assisting to pack a too coarse one. - -_Different Behavior of Coarse and Fine Ore._—In taking up a shovelful -of ore, the fine will be observed to predominate in the bottom and -center, and the coarse on the top and sides. When thrown from the -shovel, the coarse will outstrip the fine and fall beyond it. In making -a conical pile the coarse ore will roll to the base, leaving the fine -near the apex. This difference in the action of the mobile coarse ore -and the sluggish fines is the key to the practical side of feeding, -both manual and mechanical. It is not sufficient to tell the feeder to -throw the coarse in the middle and the fine against the sides; if it be -easier to do it some other way such instructions will count for little. -The desired result can be best secured by making the right way easier -than the wrong way. - -It is generally conceded that the open-top furnaces, fed by hand -through a slot in the floor-plates, do not give as satisfactory results -as the hooded furnaces with long feed-doors on both sides. In the -open-top furnace it is comparatively difficult to throw to the sides; -the narrower the slot the greater the difficulty. The major part of the -charge will drop near the center, making that place higher than the -sides. The fine ore will tend to stay where it falls, while the coarse -will tend to roll to the sides, thus leading to an arrangement of the -charge just the reverse of what it ought to be. In the hooded furnace -most of the material will naturally fall near the doors, causing the -sides to be higher than the center toward which the coarse will roll, -while the force of the throw as the ore is shoveled in will also have -a tendency to concentrate the coarse material in the center. - -Once a proper balance of conditions has been found, absolute -regularity of routine is the secret of good results. An experienced -and intelligent feeder owes his merit to his conscientious regularity -of work. He may have to vary his program somewhat when he encounters -a furnace that is suffering from the results of bad feeding by a -predecessor; but his guiding principle is first to restore regularity, -and then maintain it. A poor feeder can bring about, in a single -shift, disorders that will require many days to correct, if indeed -they are corrected at all during the campaign. The personal element is -productive of more harm than good. - -_Mechanical Feeding._—If it be admitted that the work of a feeder -is the better the more it approximates the regularity of that of a -machine, it ought to be desirable to eliminate the personal factor -entirely and design a machine for the purpose, which would be a -comparatively simple matter if it be known just what we want to -accomplish. No valid ground now exists for prejudice against mechanical -feeding in lead smelting. It is in successful operation in a number -of large works, and is being installed in others. Our furnaces have -outgrown the shovel; we have passed the limit of efficiency of the -old methods of handling material for them. We must come to mechanical -feeding in spite of ourselves. But whatever may be the motive leading -to its introduction, its chief justification will be discovered, -after it has been successfully installed and correctly adjusted, in -the consequent great improvement of general operating results, metal -saving, etc. It will remove one of the most uncertain factors with -which the metallurgist has to deal, thereby bringing into clearer view -for study and regulation the other factors (fuel and blast proportion, -slag composition, etc.) in a way that has hardly been possible under -the irregularities consequent upon hand feeding. - - - - - MECHANICAL FEEDING OF SILVER-LEAD BLAST FURNACES[12] - - BY ARTHUR S. DWIGHT - - (January 17, 1903) - - -_Historical._—A silver-lead furnace fed by means of cup and cone was in -operation in 1888 at the works of the St. Louis Smelting and Refining -Company at St. Louis, Mo., but it is probable that previous attempts -had been made, since Hahn refers (“Mineral Resources of the United -States,” 1883) in a general way to experiments with this device, which -were unsuccessful because the heat crept up in the furnace and gave -over-fire. At the time of my visit to the St. Louis works (in 1888) -the furnaces were showing signs of over-fire, but this may not have -been their characteristic condition. A. F. Schneider, who built the St. -Louis furnaces, afterward erected, at the Guggenheim works at Perth -Amboy, N. J. , round furnaces with cup and cone feeders, but although -good results are said to have been obtained, the running of refinery -products is no criterion of what they would do on general ore smelting. - -_Cup and Cone Feeders._—The cup and cone is an entirely rational device -for feeding a round furnace, but is quite unsuitable for feeding a -rectangular one. Furnaces of the latter type were installed for copper -smelting at Aguas Calientes, Mex., with two sets of circular cup and -cone feeders, but disastrous results followed the application of this -device to lead furnaces. The reason is clear when it is considered that -a circular distribution cannot possibly conform to the requirements -of a rectangular furnace. A more rational device was designed for the -works at Perth Amboy, N. J. - -[Illustration: FIG. 2.—Perth Amboy, N. J. , Lead Furnace. Vertical -section at right angles to Fig. 3.] - -_Pfort Curtain._—About ten years ago some of the American smelters -adopted the Pfort curtain, which, as adapted to their requirements, -consisted of a thimble of sheet iron hung from the iron deck plates so -as to leave about 15 in. of space between it and the furnace walls, -this space being connected with the down-take of the furnace. The -thimble was kept full of ore up to the charge-floor. This device was -popular for a time, chiefly because it prevented the furnace from -smoking and diminished the labor of feeding, but it was found to give -bad results in the furnaces, it being impossible to observe how the -charge sunk (except by dropping it below the thimble), while the -curtain had to be removed in order to bar down accretions, and, most -important, it caused irregular furnace work and high metal losses, -because it effected a distribution of the coarse and fine material -which was the reverse of correct, the evil being emphasized by the -taking off of the gases close to the furnace walls. - -[Illustration: FIG. 3.—Perth Amboy, N.J., Lead Furnace. Vertical -section at right angles to Fig. 2.] - -_Terhune Gratings._—R. H. Terhune designed a device (United States -patent No. 585,297, June 29, 1897), which comprised two grizzlies, -one on each side of the furnace, sloping downward from the edge of -the charge-floor toward the center line of the furnace. The bars -tapered toward the center of the furnace, the open spaces tapering -correspondingly toward the sides, so that as the charge was dumped on -them a classification of coarse and fine would be effected. This device -is correct in conception. - -_Pueblo System._—In the remodeling of the plant of the Pueblo Smelting -and Refining Company in 1895, under the direction of W. W. Allen, -mechanical feeding was introduced, and the system was the first one to -be applied successfully on a large scale. The furnaces of this plant -are 60 × 120 in. at the tuyeres, with six tuyeres, 4 in. in diameter -on each side, the nozzles (water cooled) projecting 6 in. inside the -jackets. The hight of the smelting column above the tuyeres is 20 ft. -The gases are taken off below the charge-floor, and the furnace tops -are closed by hinged and counter-weighted doors of heavy sheet iron, -opened by the attendant, just previous to dumping the charge-car. In -the side walls of the shaft are iron door-frames, ordinarily bricked -up, but giving access to the shaft for repairs or barring out without -interfering with the movement of the charge-car. Extending across the -shaft, about 18 in. above the normal stock line, are three A-shaped -cast-iron deflectors, dividing the area of the shaft into four equal -rectangles. - -The general arrangement of the plant is shown in Fig. 4. From the -charge-car pit there extends an inclined trestle, on an angle of 17 -deg. to the charge-floor level, in line with the battery of furnaces. -The gage of the track is approximately equal to the length of the -furnaces at the top. The charge-car, actuated by a steel tail-rope, -moves sideways on this track from the charging-pit to any furnace -in the battery. The hoisting drums are located at the crest of the -incline, inside of the furnace building. At the far end of the latter -there is a tightener sheave, with a weight to keep proper tension on -the tail-rope. The charge-car has a capacity of 5 tons. It has an -A-shape bottom, and is so arranged that one attendant can quickly trip -the bolt and discharge the car. - -[Illustration: FIG. 4.—Pueblo System. Longitudinal vertical section -through incline.] - -While the car is making its trip the charge-wheelers are filling their -buggies, working in pairs, each man weighing up a halfcharge of a -particular ingredient. They then separate, each taking his proper place -in the line of wheelers on either side. When the car has returned, the -wheelers successively discharge their buggies into opposite ends of -the car. The coke is added last, to avoid crushing. The system is not -strictly economical of labor, since the wheelers, who must always be -ready for their car, have to wait for its return, which necessitates -more wheelers than would otherwise be required. Figs. 5, 6 and 7 show -the car. - -[Illustration: FIG. 5.—Pueblo Charge-car. (Side elevation.)] - -A vertical section through the car filled by dumping from the two ends -will show an arrangement of coarse and fine, which is far from regular. -Analyzing its structure, we shall find a conical pile near each end, -with a valley between them, in which coarse ore will predominate. The -deflectors in the furnace, previously referred to, serve to scatter -the fines as the charge is dropped in. Without them the feeding of the -furnace would be a failure; with them it is successful, though not so -completely as might be, the furnaces having a tendency to run with hot -tops. With the battery of seven furnaces, each smelting an average of -100 tons of ore per day, the saving, as compared with hand-feeding, -was $63 per day, or 9c. per ton of ore, this including cost of steam, -but not wear and tear on the machinery. This is distinctly a maximum -figure; with fewer furnaces the fixed charges of the mechanical feed -would soon increase the cost per ton to such a figure that the two -systems would be about equal in economy. - -[Illustration: FIG. 6.—Pueblo Charge-car. (Plan.)] - -[Illustration: FIG. 7.—Pueblo Charge-car. (End elevation.)] - -_East Helena System._—This was introduced at the East Helena plant of -the United Smelting and Refining Company by H. W. Hixon. The plant -comprised four lead furnaces, each 48 × 136 in., with a 21 ft. smelting -column. They were all open-top furnaces, fed through a slot over the -center, the gases being taken off below the floor. They were capable of -smelting about 180 tons of charge (ore and flux) per 24 hours, using -a blast of 30 to 48 oz., furnished by two Allis duplex, horizontal, -piston blowers, air-cylinders 36 in. diam., 42 in. stroke, belted -from electric motors. The Hixon feed was designed to meet existing -conditions, without irrevocably cutting off convenient return to -hand feeding in case of an emergency. As shown in Fig. 9 there is a -track-way at right angles to the line of furnaces. The car hoisted up -the incline is landed on a transfer carriage, on which, after detaching -the cable, it can be moved over the tops of the furnaces by means of -a tail-rope system. The gage of the charge-car is 4 ft. 9 in.; of -the transfer carriage, 11 ft. 8 in. A switch at the lower end of the -incline permits two charge-cars to be employed, one being filled while -the other is making the trip. In sending down the empty car a hand -winch is necessary to start it from the transfer carriage. Figs. 10 and -11 show the charge-car; Fig. 12 the transfer carriage. - -[Illustration: FIG. 8.—Pueblo System. (Sectional diagrams of furnace -top.)] - -The charge-car is 10 × 4 × 3.5 ft., and has capacity for 6 tons of ore, -flux, slag and fuel, the total of ore and flux being usually 8800 lb. -Its bottom is flat, consisting of two doors, hinged along the sides -and kept closed by means of chains wound about a longitudinal windlass -on top of the car. The charging pits are decked with iron plates, -leaving a slot along the center of each car exactly like the slot in -the furnace top. The loaded ore-buggies are taken from the wheelers by -two men, who carefully distribute the contents of each buggy along the -whole length of the charge-car by dragging it along the slot while in -the act of dumping. Each buggy contains but one ingredient; they follow -one another in a prescribed order, so as to secure thin layers in the -charge-car. The coke is divided into three or more layers. - -[Illustration: FIG. 9.—East Helena System. (Vert-longitudinal section -and plan of incline.)] - -[Illustration: FIG. 10.—East Helena Charge-car. (Side elevation.)] - -The first few trials of this device were not satisfactory. The furnaces -quickly showed over-fire, and decreased lead output, which would not -yield to any remedy except a return to hand feeding. The total charge -being dropped in the center of the furnace, a central core of fines -was produced, the lumps tending to roll toward the walls. This wrong -tendency was emphasized by the presence of the chains supporting -the bottom of the charge-car. On unwinding them to dump the car, -the doors were prevented from dropping by the wedging of the chains -in the charge, which in turn arched itself more or less against the -sides of the car; hence the doors opened but slowly, and often had to -be assisted by an attendant with a bar. In consequence of this slow -opening, considerable fine ore sifted out first and formed a ridge in -the center of the furnace, from the slopes of which the coarser part of -the charge, the last to fall, naturally rolled toward the sides. This -fact, determined during a visit of the writer in April, 1899, proved -to be the key to the situation. The attendant operating the tail-rope -mechanism was instructed to move the transfer carriage rapidly backward -and forward over the slot while the first one-third or one-half of -the charge was dropping, and during the rest of the discharge to let -the car stand directly over the slot and permit the coarser material -to fall in the center of the furnace. Two piles of comparatively fine -material were thus left on the charge-floor, one on each side of the -slot. These were subsequently fed in by hand, with instructions to -throw the material well to the sides of the furnace. - -[Illustration: FIG. 11.—East Helena Charge-car. (Plan.)] - -The furnaces were running very hot on top when this modified procedure -was begun. In a few hours the over-fire had disappeared; the lead -output was increasing; and the furnaces were running normally. This was -done about May 1, 1899, and from that time until about February 20, -1900, the Hixon feed, as modified above, was continuously in operation. -In October, 1898, with three furnaces in operation and hand feeding, -the labor cost per furnace was $42.06 per day; in October, 1899, with -the same number of furnaces and mechanical feeding, it was $41 per day, -the saving being only 0.6c. per ton of charge. - -[Illustration: FIG. 12.—East Helena Charge-car and Transfer Carriage. -(Elevation.)] - -[Illustration: FIG. 13.—East Helena System, with spreader and curtains. -(Experimental form.)] - -_Dwight Spreader and Curtain._—In January, 1900, the writer again -had occasion to visit the East Helena plant, to investigate why a -certain cheap local coke could not be used successfully instead of -expensive Eastern coke. Strange as it may seem, the peculiar behavior -of the cokes was traced to improper feeding of the furnaces. Further -study of the mechanical feeding system, then in operation for nine -months, showed that it was far from perfect, and it appeared desirable -to design a spreader which would properly distribute the material -discharged from the Hixon car and dispense with hand feeding entirely. -An experimental construction was arranged, as shown in Fig. 13. The -flanged cast-iron plates around the feeding slot were pushed back and -a roof-shaped spreader, with slopes of 45 deg., was set in the gap, -leaving openings about 8 in. wide on each side. The plan provided for -two iron curtains to be hung, one on each side of the spreader, and so -adjusted that the fine ore sliding down the spreader would clear the -edge of the curtain and shoot toward the sides of the furnace, while -the coarse ore would strike the curtain and rebound toward the center -of the furnace. The classification effected in this manner was capable -of adjustment by raising or lowering the curtain. This arrangement was -found to work surprisingly well. The first furnace equipped with it -immediately showed improvement. It averaged better in speed, with lower -blast, lower lead in slag and matte, and better bullion output than -the other furnaces operating under the old system. The success of the -spreader and curtain being established, the furnaces were provided with -permanent constructions, the only modifications being that the ridge of -the spreader was lowered to correspond with the level of the floor and -the curtains were omitted, the feeding being apparently satisfactory -without their aid. In their absence, the lowering of the spreader was a -proper step, as it distributed the material fully as well, and caused -less abrasion of the walls. The final form is shown approximately -in Fig. 14. It has given complete satisfaction at East Helena since -February, 1900, and has been adopted as the basis for the mechanical -feeding device in the new plant of the American Smelting and Refining -Company at Salt Lake, Utah. - -[Illustration: FIG. 14.—East Helena System. (Final form, approximate.)] - -_Comparison of Systems._—In mechanical design the Pueblo system -is better than the East Helena, being simpler in construction and -operation. No time is lost in attaching and changing cables, operating -transfer carriage, etc. In both systems the track runs directly over -the tops of the furnaces, and this is an inconvenience when furnace -repairs are under way. The Pueblo car is the simpler, and makes the -round trip in about half the time of a car at East Helena, so the two -cars of the latter do not make much difference in this respect. The -system of filling the charge-car at Pueblo is also the quicker. It may -be estimated roughly that per ton of capacity it takes 2.5 to 3 times -as long to fill the East Helena car; and this means longer waiting on -the part of the wheelers, and consequently greater cost of moving the -material, representing probably 7 or 8c., in favor of Pueblo, per ton -of charge handled. However, both systems are wasteful of labor. As to -furnace results, it is believed that the better distribution of the -charge in the East Helena system leads to greatly increased regularity -of furnace running, less tendency to over-fire, some economy in fuel, -less accretions on the furnace walls and larger metal savings. If the -half of these conclusions are true, the difference of 7 or 8c. per ton -in favor of the Pueblo system, which can be traced almost entirely -to the cost of filling the charge-car, sinks into insignificance -in comparison with the important advantages of having the furnaces -uniformly and correctly fed. - -_True Function of the Charge-Car._—The radically essential feature of a -mechanical feeding device is that part which automatically distributes -the material in the furnace, whatever approximate means may have been -used to effect the delivery. - -Taking a hasty review of the numerous feeding devices that have -been tried in lead-smelting practice, we cannot but remark the fact -that those which depended upon dumping the charge into the furnace -from small buggies or barrows failed generally to secure a proper -classification and distribution of coarse and fine, and, consequently, -were abandoned as unsuccessful, while the adoption of the idea of the -charge-car for transporting the material to the furnace in large units -seems to have been coincident with a successful outcome. It is natural -enough, therefore, that the car should be regarded by many as the vital -feature. This view of the question is not, however, in accordance -with the true perspective of the facts, and merely limits the field -of application in an entirely unnecessary way. It must be apparent -that the essential function of the charge-car is cheap and convenient -transportation. The distribution of the charge is an entirely different -matter, in which, however, the charge-car may be made to assist, as -in the Pueblo system; or entirely distinct and special means may be -employed for the distribution, as in the East Helena system. - -To follow the argument to its conclusion, let us imagine for the moment -that the East Helena plant were arranged on the terrace system, with -the furnace tops on a level with the floor of the ore-bins. Certain -precautions being observed, the spreader would give as good results -with small units of charge delivered by buggies as it now does with the -large units delivered by the charge-car, and the expense of delivery -to the furnaces would be practically no more than it now is to the -charge-car pit. The furnace top would, of course, have to be arranged -so that the buggies, in discharging, could be drawn along the slot, -so as to give the necessary longitudinal distribution parallel to the -furnace walls, just as is now done in filling the charge-car. The ends -of the spreader, if built like a hipped roof, would secure proper -feeding of the front and back. - -Thus, by eliminating the charge-car, and with it the necessity for -powerful hoisting machinery, with its expensive repairs and operating -costs, we may greatly simplify the problem of mechanical feeding, and -open the way for the adoption of successful automatic feeding in many -existing plants where it is now considered impracticable. - - - - - COST OF SMELTING AND REFINING - - BY MALVERN W. ILES - - (August 18, 1900) - - -In the technical literature of lead smelting there is a lamentable lack -of data on the subject of costs. The majority of writers consider that -they have fulfilled their duties if they discuss in full detail the -chemical and engineering sides of the subject, leaving the industrial -consideration of cost to be wrought out by experience. When an engineer -or metallurgist collects data on the costs involved in the various -smelting operations, he generally hesitates to give this special -information to the public, as he regards it as private, or reserves it -as stock in trade to be held for his own use. - -The following tables of cost have been compiled from actual results -of smelting and refining at the Globe works, Denver, Colo., and are -offered in the hope that they will prove a valuable addition to the -literature of lead smelting. These results are offered tentatively, -and, while true for the periods stated, they require considerable -adjustment to meet the smelting conditions of the present time. - - -COST OF HAND-ROASTING PER TON (2000 LB.) OF ORE - - 1887 $3.975 - 1888 4.280 - 1889 4.120 - 1890 3.531 - 1891 3.530 - 1892 - 1893 - 1894 3.429 - 1895 2.806 - 1896 2.840 - 1897 2.740 - 1898 2.620 - -At first the roasting was done mainly by hand roasters; later two -Brown-O’Harra mechanical furnaces were used, and the cost was reduced, -but not to the extent usually conceded to this type of furnace, as the -large amount of repairs and the consequent loss of time diminished -the apparent gain due to greater output. The figures quoted above may -be considered somewhat higher than the average, as the roasters were -charged in proportion with expenses of general management, office, etc. - -In viewing the yearly reduction of costs one must take into -consideration many changes in the furnace construction and working, as -well as the items of labor, fuel, etc. From 1887 to 1899 the principal -changes in the construction of the hand-roasting furnaces consisted -in an increase of width, 2 ft., which allowed an addition of 200 lb. -to each ore charge, and corresponded to a total increase per furnace -of 1200 lb. in 24 hours. In the working of the charge an important -change was made in the condition of the product. Formerly the material -was fused in the fusion-box and drawn from the furnace in a fused or -slagged condition; and while this gave an excellent material for the -subsequent treatment in the shaft furnace in that there was very little -dusting of the charge, and a considerable increase in the output of the -furnace, the disadvantages of large losses of lead and silver greatly -over-balanced the advantages, and called for an entire abandonment of -the fusion-box. As a result of experience it was found that the best -condition of product is a semi-fused or sintered state, in which the -particles of roasted ore have been compressed by pounding the material, -which has been drawn into the slag pots, with a heavy iron disk. The -amount of “fines” under these conditions is quite small and depends -upon the percentage of lead in the ore, the degree of heat employed, -and the extent of the compression. - -The total cost was partly reduced from the lessened labor cost -following the financial disturbance of 1893, and partly from the -reduction in the fuel cost, the former expensive lump coal being -replaced by the slack coals from southern Colorado. - -The comparison of the cost of labor by the two methods shows a gain of -54c. a ton in favor of the mechanical furnaces. However, I consider -that this gain is a costly one, and is more than offset by the large -amount of high-grade fuel required, and the expense of repairs not -shown in the following table. Indeed, I believe that at the end of five -or ten years the average cost of roasting per ton by the hand roasters -will be even smaller than by these mechanical roasters. - -To illustrate the details of roasting cost and to furnish a comparison -of the hand roasters and mechanical furnaces, the following table has -been prepared: - - DETAILS OF AVERAGE MONTHLY COST FOR 1898 OF HAND ROASTERS AND - MECHANICAL FURNACES - - ───────────┬────────┬───────┬─────────────────────┬───────────────────── - │ │ │ HAND ROASTERS │ BROWN-O’HARRA - │ │ │ │ MECHANICAL FURNACES - │ TOTAL │ TONS ├──────┬──────┬───────┼──────┬──────┬─────── - Month │ TONS │ROASTED│LABOR │ COAL │GENERAL│LABOR │COAL │GENERAL - │ROASTED │PER DAY│ $ │ $ │EXPENSE│ $ │ $ │EXPENSE - │ │ │ │ │ $ │ │ │ $ - ───────────┼────────┼───────┼──────┼──────┼───────┼──────┼──────┼─────── - January │ 5,691 │ 184 │ 1.47 │ 0.53 │ 0.80 │ 0.92 │ 0.80 │ 1.32 - February │ 5,677 │ 203 │ 1.44 │ 0.44 │ 0.99 │ 0.72 │ 0.58 │ 1.01 - March │ 5,821 │ 188 │ 1.51 │ 0.53 │ 0.64 │ 0.76 │ 0.64 │ 0.62 - April │ 5,472 │ 182 │ 1.47 │ 0.47 │ 0.71 │ 0.80 │ 0.69 │ 0.87 - May │ 5,444 │ 176 │ 1.55 │ 0.51 │ 0.84 │ 0.80 │ 0.69 │ 0.81 - June │ 4,859 │ 162 │ 1.58 │ 0.48 │ 0.71 │ 0.90 │ 0.68 │ 1.17 - July │ 5,691 │ 184 │ 1.59 │ 0.48 │ 0.75 │ 0.72 │ 0.56 │ 0.64 - August │ 5,910 │ 191 │ 1.55 │ 0.46 │ 0.83 │ 0.72 │ 0.55 │ 0.75 - September │ 5,677 │ 189 │ 1.55 │ 0.45 │ 0.74 │ 0.73 │ 0.55 │ 0.67 - October │ 6,254 │ 202 │ 1.48 │ 0.49 │ 0.72 │ 0.65 │ 0.50 │ 0.60 - November │ 6,291 │ 213 │ 1.42 │ 0.47 │ 0.80 │ 0.66 │ 0.53 │ 0.70 - December │ 5,874 │ 198 │ 1.45 │ 0.48 │ 0.78 │ 0.79 │ 0.63 │ 0.81 - ├────────┼───────┼──────┼──────┼───────┼──────┼──────┼─────── - Average │ │ │ 1.50 │ 0.48 │ 0.77 │ 0.76 │ 0.62 │ 0.83 - Total │ │ │ │ │ 2.75 │ │ │ 2.21 - ───────────┴────────┴───────┴──────┴──────┴───────┴──────┴──────┴─────── - -_Cost of Smelting._—The lead-ore mixtures of the United States, in -addition to lead, contain gold, silver and generally copper, and are -treated to save these metals. The total cost of smelting is made up of -a large number of items. The questions of locality and transportation, -fuel, fluxes and labor are the principal factors, to which must be -added the handling of the material to and from the furnace; the -furnace itself, its size, shape, and method of smelting, the volume -and pressure of blast, etc. The following table of costs, from 1887 to -1898, shows in a general way the great advance that has been made in -the development of smelting, and the consequent reduction in cost per -ton of ore treated: - - -AVERAGE COST OF SMELTING, PER TON - - 1887 $4.644 - 1888 4.530 - 1889 4.480 - 1890 4.374 - 1891 4.170 - 1892 4.906 - 1893 3.375 - 1894 3.029 - 1895 2.786 - 1896 2.750 - 1897 2.520 - 1898 2.260 - -In connection with this table of smelting cost should be considered the -changes developed during the interval 1887-1889, outlined as follows: - -CONDITIONS OF SMELTING IN 1886 AND 1899 CONTRASTED TO SHOW THE PROGRESS - OF DEVELOPMENT - - ────┬───────────────┬────────────┬───────────────┬─────────────┐ - │AREA OF FURNACE│ HEIGHT OF │BLAST PRESSURE,│ FORE HEARTH │ - │ AT TUYERES, │CHARGE FROM │ LB. PER │CAPACITY, CU.│ - │ IN. │TUYERES, FT.│ SQ. IN. │ FT. │ - ────┼───────────────┼────────────┼───────────────┼─────────────┤ - 1886│ 30 × 100 │ 11 │ 1 │ 6 │ - │ │ │ │ │ - │ │ │ │ │ - 1899│ 42 × 140 │ 16 │ 3 to 4 │ 128 │ - │ │ │ │ │ - ────┴───────────────┴────────────┴───────────────┴─────────────┘ - - ────┬────────────┬────────┬───────────────┬───────────────┐ - │ SLAG │ FUEL │ SLAG REMOVED, │MATTE REMOVED, │ - │ SETTLED │ │ LB. PER TRIP │ LB. PER │ - ────┼────────────┼────────┼───────────────┼───────────────┤ - 1886│ │ │ │ TRIP │ - │ In pots │Charcoal│ By hand │ By hand │ - │ │ │ 280 │ 200 │ - 1899│ │ │ │ │ - │In furnaces │ Coke │ By locomotive │ By horse │ - │ │ │ 3000-6000 │ 2000-3000 │ - ────┴────────────┴────────┴───────────────┴───────────────┘ - -I believe that there is room for further improvement in the -substitution of mechanical transportation within the works for hand -labor, and that the fuel cost can be materially reduced by replacing -the coke, which at present contains 16 to 22 per cent. of ash, by a -fuel of purer and better quality. - -_Cost of Refining by the Parkes Process._—In general it may be stated -that the average cost of refining base bullion is from $3 to $5 a -ton. This amount is based on the cost of labor, spelter, coal, coke, -supplies, repairs and general expenses. When the additional items -of interest, expressage, brokerage and treatment of by-products are -considered, which go to make up the total refining cost, the amount may -be stated approximately as $10 per ton of bullion treated. - -Variations in the cost occur from time to time, and are due to several -causes, principally the irregularity of the bullion supply and its -consequent effect on the work of the plant. When the amount of bullion -available for treatment is small, the plant cannot be run to its -maximum capacity, and the cost per ton will naturally be increased. To -illustrate this variation, the average cost per ton of base bullion -refined during nine months in 1893 was: - -January, $4.864; February, $5.789; March, $5.024; April, $3.915; May, -$5.094; June, $4.168; July, $4.231; August, $4.216; September, $5.299. - -The yearly variation shows but little change, as the average cost per -ton was for 1893, $4.75; for 1894, $3.99; for 1895, $4.21; for 1896, -$3.90. In considering the total cost of refining, the additional -factors of interest, expressage, parting, brokerage, and reworking of -by-products must be considered. As the doré silver is treated at the -works or elsewhere, so will the total cost be less or greater. The -following table gives the cost in detail, when the parting is done at -the same works: - - - AVERAGE MONTHLY COST OF REFINING PER TON OF BULLION TREATED - - ─────────────────────┬────────────┬────────────┬────────────┬───────── - ITEMS │ 1895 │ 1895 │ 1896 │ AVERAGE - │JAN. TO JULY│JULY TO DEC.│JAN. TO JULY│ - ─────────────────────┼────────────┼────────────┼────────────┼───────── - Labor │ $2.351 │ $1.718 │ $1.836 │ $1.968 - Spelter │ 0.757 │ 0.840 │ 0.987 │ 0.861 - Coal │ 0.585 │ 0.442 │ 0.461 │ 0.496 - Coke │ 0.634 │ 0.418 │ 0.511 │ 0.521 - Supplies, repairs and│ │ │ │ - general expenses │ 0.343 │ 0.273 │ 0.252 │ 0.289 - Interest │ 1.808 │ 1.075 │ 1.070 │ 1.317 - Expressage │ 1.360 │ 1.015 │ 0.882 │ 1.085 - Parting and brokerage│ 2.483 │ 2.084 │ 1.796 │ 2.121 - Reworking by-products│ 1.567 │ 1.286 │ 1.625 │ 1.492 - ├────────────┼────────────┼────────────┼───────── - Totals │ $11.888 │ $9.151 │ $9.420 │ $10.151 - Tons bullion refined │5,511.58 │9,249.07 │10,103.43 │8,287.99 - ─────────────────────┴────────────┴────────────┴────────────┴───────── - - -An analysis of the different items of cost is important, and a brief -summary is given below. - -_Labor and Attendance._—The cost for this item varies but little from -year to year, and its reduction depends, for the most part, on a larger -yield per man rather than on a reduction of wages. If a man at the same -or slightly increased cost can give a larger output, so will the labor -cost per ton be diminished. This result is accomplished by enlarging -the furnace capacity and by using appliances which will handle the -bullion and its products in an easier and quicker manner. The small -size of the furnaces, settlers and retorts used at modern refineries is -open to criticism; I believe that great improvement can be made in this -direction. - -_Spelter._—The cost of this item varies with the market conditions, -and will probably be changed but little in the future, as the amount -necessary per ton of bullion seems to be fixed. - -_Coal._—The amount required per ton of bullion is fairly constant, and -while lessened cost for fuel may be attained by the substitution of oil -or gaseous fuel, the fuel cost in comparison with the aggregate cost is -very small, and leaves little opportunity for improvement in this line. - -_Supplies._—This item includes brooms, shovels, wheelbarrows, etc., and -the amount is small and fairly constant from year to year. - -_Repairs._—This item is quite small in works properly constructed; -and in this connection I wish to call particular attention to the -floor covering, which should be made of cast-iron plates from 1.5 to -2 in. thick, and placed on a 2 to 3 in. layer of sand spread over the -well-tamped and leveled ground. The constant patching of brick floors -is not only an annoyance, but is costly from the additional labor -required. Furthermore, a brick floor does not permit a close saving of -the metallic scrap material. - -It will be found economical in the long run to protect all exposed -brickwork of furnaces or kettles with sheet iron. - -In the construction of the refinery building I should advise brick -walls except at the end or side, where there is the greatest likelihood -of future extension; here corrugated iron may be used. The roof should -not be made of corrugated iron, as condensed or leakage water is liable -to collect and drop on those places where water should be scrupulously -avoided. The presence of water in a mold at the time of casting, even -though small in amount, will cause explosions and will scatter the -molten lead, endangering the workmen. - -The item of repair for the ordinary corrugated iron roof may be -diminished by constructing it of 1 in. boards with intervening spaces -of half an inch, the whole overlaid with tarred felt, and covered with -sheets of iron at least No. 27 B. W. G., painted with graphite paint -and joined together with parallel rows of ribbed crimped iron. - -_General Expenses._—This item is generally constant, and calls for no -special comment. - -_Interest._—This important item is, as a rule, considerable, as the -stock of bullion and other gold-and silver-bearing material is quite -large. For this reason special attention should be given to prevent -the accumulation of stock or by-products. The occasional necessity of -additional capital to run the business should preferably be met by an -increase of working capital, rather than by a direct loan. - -_Expressage._—This item, as a rule, is large, and should be taken into -consideration in the original plans for the location of the refining -works. - -_Parting._—The item of parting and brokerage is the largest of the -refinery costs, and for obvious reasons a modern smelting plant should -have a parting plant under its own control. - -_The Working of the By-Products._—This constitutes a large item of -cost, and considerable attention should be devoted to the improvement -of present methods, which seem faulty, slow and expensive. - -_Summary._—The items of smaller cost with their respective amounts per -ton of base bullion treated are: Spelter, $0.85; coal, $0.50; coke, -$0.50; supplies, repairs and general expenses, $0.35; total, $2.10. It -is doubtful whether much improvement can be made in the reduction of -these costs. - -The items of larger cost are: Labor, $2; interest, $1.32; expressage, -$1.10; parting and brokerage, $2; reworking by-products, $1.50; total, -$7.92. The general manager usually attends to the items of interest, -expressage and brokerage, leaving the questions of labor and working of -by-products to the metallurgist. - -The cost quoted for smelting practice, as employed at Denver, will -differ necessarily from those at other localities, where the cost of -labor, freight rates on spelter, fuel, etc., are changed. Refining -can doubtless be done at a lower cost at points along the Mississippi -River, and even more so at cities on the Atlantic seaboard, as Newark -or Perth Amboy, N. J. - -The consolidation of many of the more important smelting plants of the -United States under one management will doubtless alter the figures of -cost given above, particularly as the interest cost there stated is at -the high rate of 10 per cent., a condition of affairs now changed to 5 -per cent. Other factors have lessened the cost of refining; the bullion -produced at the present time is softer, or contains a smaller amount -of impurities, and admits of easier working with shorter time and less -labor. By proper management larger tonnages are turned out per man, and -the Howard stirrer and Howard press have simplified and cheapened the -working of the zinc skimmings. To illustrate the comparatively recent -conditions of cost I have compiled the following table for each month -of the year 1898: - - -COST OF REFINING DURING 1898, INCLUDING LABOR, SPELTER, COAL, COKE, -SUPPLIES, REPAIRS AND GENERAL EXPENSES. - - January $3.59 - February 3.28 - March 3.26 - April 3.59 - May 3.38 - June 3.56 - July 3.65 - August 3.54 - September 3.35 - October 3.45 - November 3.20 - December 3.56 - Average cost during the year, $3.45. - -It is understood, of course, that these figures do not include cost of -interest, expressage, parting, brokerage and reworking of by-products. - - [Although this article refers to conditions in 1898, since which time - there have been improvements in practice, the latter have not been of - radical character and the figures given are fairly representative of - present conditions.—EDITOR.] - - - - - SMELTING ZINC RETORT RESIDUES[13] - - BY E. M. JOHNSON - - (March 22, 1906) - - -The following notes were taken from work done at the Cherokee -Lanyon Smelter Company, Gas, Kansas, in 1903. It was practically an -experiment. The furnace was only 36 × 90 in. at the crucible, with a -10 in. side bosh and a 6 in. end bosh. There were five tuyeres on each -side with a 3 in. opening. The side jackets measured 4.5 ft. × 18 in. -The distance from top of crucible to center of tuyeres was 11.5 in. - -The blast was furnished by one No. 4½ Connellsville blower. The -furnace originally was only 11 ft. from the center of tuyeres to the -feed-floor, and had only been saving about 60 per cent. of the lead. -This loss of lead, however, was not entirely due to the low furnace. -As no provision had been made to separate the slag and matte, upon -assuming charge I raised the feed-floor 3 ft., thereby changing the -distance from the tuyere to top of furnace from 11 ft. to 14 ft. Matte -settlers were also installed. These two changes raised the percentage -of lead saved to 92, as shown by monthly statements. The furnace being -small, and a high percentage of zinc oxide on the charge, the campaigns -were naturally short. The longest run was about six weeks. This was -made on some residue that had been screened from the coarse coal, and -coke, and had weathered for several months. This particular residue -also carried about 10 per cent. lead. The more recent residue that had -not been screened and weathered, and was low in lead, did not work so -well. Although these residues consisted of a large proportion of coal -and coke, it seemed impossible to reduce the percentage of good lump -coke on the charge lower than 12.5 or 13 per cent. At the same time the -reducing power of the residue was strong, and with the normal amount of -coke caused some trouble in the crucible. - -When residue containing semi-anthracite coal was smelted, the saving -in lead dropped, and the fire went to the top of the furnace, burning -with a blue flame, thereby necessitating the reduction of this class of -material. This residue had been screened through a five-mesh screen, -and wet down in layers, becoming so hard that it had to be blasted. -The low saving of lead with this class of material was a surprise, -as it has been claimed that the substitution of part of the fuel by -anthracite coal did not affect the metallurgical operations of the -furnace. - -The slag was quite liquid and flowed very well at all times. However, -there was a marked variation in the amount at different tappings. This, -I am satisfied, was not due to irregular work on the furnace, but may -be accounted for in the following manner. The residue (not screened or -weathered to any extent), consisting approximately of one-half coal -and coke, was very bulky, and while there was about 35 per cent. of it -on the charge by weight, there was over 50 per cent. of it by bulk, -not including slag and coke. In feeding, therefore, it was a difficult -matter to mix the whole of it with the charge. Several different ways -of feeding the furnace were tried. The one giving the most satisfactory -results was to feed nearly all of the residue along the center of the -furnace, in connection with the lime-rock, coarse ore and coarse iron -ore, and the fine and easy smelting ores along the sides. The slag was -spread uniformly over the whole furnace, while the sides were favored -with the coke. The charge would drop several inches at a time, going -down a little faster in the center than on the sides. - -It is possible that a small proportion of the residue in connection -with the easy smelting, leady, neutral ore, iron ore and lime-rock -formed the type of slag marked No. 1. - - ───┬───────┬──────┬─────┬──────┬─────┬─────┬──── - │ SiO₂ │ FeO │ MnO │ CaO │ ZnO │ Pb │ Ag - ───┼───────┼──────┼─────┼──────┼─────┼─────┼──── - 1 │ 33.7 │ 34.1 │ 1.0 │ 16.5 │ 7.5 │ 0.9 │ 0.7 - 2 │ 31.0 │ 36.1 │ 1.2 │ 16.0 │ 9.6 │ 1.3 │ - ───┴───────┴──────┴─────┴──────┴─────┴─────┴──── - -This being tapped with a good flow of slag, the charge would drop, -bringing a proportionately large amount of residue in the fusion zone -which formed the type of slag marked No. 2. There was also a marked -variation in the slag-shells from different pots. The above cited -irregularities of course exist to a certain extent in any blast furnace. - - - AVERAGE ANALYSIS OF MATERIALS SMELTED - - NAME ROW NAME ROW - - Mo. iron ore A Roasted matte[15] F - Lime rock B Barrings G - Mo. galena C Coke ash H - Av. of beds D Coke[16] J - Residue[14] E - - ────┬──────┬─────┬────┬────┬────┬─────┬─────┬────┬────┬───┬────┬──── - │ SiO₂ │ FeO │CaO │MgO │ZnO │Al₂O₃ │Fe₂O₃ │ S │ Pb │Cu │ Ag │ Au - ────┼──────┼─────┼────┼────┼────┼─────┼─────┼────┼────┼───┼────┼──── - A │ 10.0 │ 65.0│ │ │ │ │ │ │ │ │ │ - B │ 1.5 │ │52.0│ │ │ │ │ │ │ │ │ - C │ 1.5 │ 2.4│ │ │ 9.5│ │ │11.0│74.0│ │ │ - D │ 50.8 │ 16.2│ │ │ 4.6│ │ │ 3.3│ 9.1│ │ │ - E │ 10.5 │ 38.5│ │ │18.0│ │ │ 4.8│ 2.2│1.0│10.0│0.03 - F │ 9.0 │ 48.0│ 3.0│ │10.0│ │ │ 4.0│ 9.9│3.0│21.0│0.06 - G │ 18.8 │ 24.4│ 5.0│ │14.5│ │ │ 6.0│25.4│ │13.0│0.07 - H │ 27.0 │ │14.9│ 4.5│ │ 19.7│ 31.6│ │ │ │ │ - │ H₂O │ V.M.│F.C.│ Ash│ S │ │ │ │ │ │ │ - J │ 1.2 │ 2.3 │85.7│11.1│ 0.9│ │ │ │ │ │ │ - ─────┴──────┴─────┴────┴────┴────┴─────┴─────┴────┴────┴───┴────┴──── - - - ANALYSIS OF BULLION, SLAG AND MATTE PRODUCED - - │/-BULLION-\ /—————————————SLAG———————————————-\/————-MATTE————-\ - │ Ag │ Au │SiO₂ │FeO │MnO│CaO │ZnO │ Pb │ Ag │ Ag │ Au │Pb │Cu - ├──────┼────┼─────┼────┼───┼────┼────┼────┼────┼────┼────┼───┼─── - Feb. │ 90.0 │1.15│31.2 │35.9│1.0│14.5│10.3│0.88│0.98│19.0│0.04│8.7│1.5 - March│ 93.1 │1.63│31.3 │37.2│1.0│13.9│11.1│0.71│1.30│21.0│0.06│8.0│2.5 - April│104.3 │1.59│29.8 │37.7│2.7│13.9│11.4│0.52│1.40│23.0│0.07│7.0│3.5 - May │ 90.0 │1.24│30.0 │37.3│2.2│14.1│ 9.3│0.86│1.10│25.4│0.07│5.1│4.0 - July │ 78.7 │1.00│32.2 │37.4│1.0│13.9│ 9.8│0.50│1.15│21.3│0.03│8.9│4.0 - Aug. │ 90.8 │1.21│31.2 │37.1 1.7│13.7│ 9.6│1.10│1.60│23.1│0.08│9.8│3.0 - Sept.│ 65.3 │2.58│32.0 │39.7│0.8│14.1│ 8.1│0.80│1.30│18.6│0.06│7.6│2.3 - ├──────┼────┼─────┼────┼───┼────┼────┼────┼────┼────┼────┼───┼─── - Avge.│ 87.5 │1.49│31.1 │37.5│1.5│14.1│10.0│0.77│1.26│21.6│0.06│7.8│3.0 - ─────┴──────┴────┴─────┴────┴───┴────┴────┴────┴────┴────┴────┴───┴─── - - - MONTHLY RECORD OF FURNACE OPERATIONS - - ─────────┬──────┬───────┬─────────┬─────────┬─────────┬─────────┐ - │BLAST │ TONS │PER CENT.│PER CENT.│PER CENT.│PER CENT.│ - │OUNCES│ PER │ PB. ON │ COKE ON │ SLAG ON │ S ON │ - │ │ F.D. │ CHARGE │ CHARGE │ CHARGE │ CHARGE │ - ─────────┼──────┼───────┼─────────┼─────────┼─────────┼─────────┤ - Feb. │ 21 │ 42.5 │ 9.0 │ 12.0 │ 30.0 │ 3.7 │ - March │ 21 │ 44.8 │ 9.7 │ 13.5 │ 37.0 │ 4.0 │ - April │ 21 │ 43.7 │ 9.0 │ 13.5 │ 35.0 │ 4.3 │ - May │ 21 │ 49.4 │ 10.0 │ 13.5 │ 30.0 │ 3.5 │ - July │ 17 │ 41.0 │ 9.8 │ 12.5 │ 34.0 │ 3.8 │ - August │ 18 │ 47.0 │ 9.3 │ 13.0 │ 32.0 │ 3.7 │ - Sept.[17]│ 15 │ 51.0 │ 7.3 │ 13.0 │ 30.0 │ 2.8 │ - ─────────┼──────┼───────┼─────────┼─────────┼─────────┼─────────┤ - Average │ │ 45.6 │ 9.1 │ 13.0 │ 32.6 │ 3.7 │ - ─────────┴──────┴───────┴─────────┴─────────┴─────────┴─────────┘ - - ────────┬────────┬─────────────────────┐ - │ MATTE │ SAVING │ - │PRODUCED│ AG AU PB │ - ────────┼────────┼──────┬───────┬──────┤ - Feb. │ 8.0} │ 84.4 │ 83.0 │ 90.3 │ - March │ 9.0} │ │ │ │ - April │ 10.0 │ 97.9 │ 70.5 │ 96.6 │ - May │ 6.5 │ 95.6 │ 109.5 │ 88.8 │ - July │ 6.0 │ 97.9 │ 90.0 │ 92.9 │ - August │ 6.3 │ 86.2 │ 107.5 │ 87.6 │ - Sept. │ 4.6 │ 92.9 │ 94.0 │ 95.6 │ - ────────┼────────┼──────┼───────┼──────┤ - Average │ 7.2 │ 90.8 │ 92.4 │ 92.0 │ - ────────┴────────┴──────┴───────┴──────┘ - -I believe that, in smelting residues high in zinc oxide, better -metallurgical results would be obtained by using a dry silicious ore in -connection with a high-grade galena ore, provided the residue be low in -sulphur. This was confirmed to a certain degree in actual practice, as -the furnace worked very well upon increasing the percentage of Cripple -Creek ore on the charge. This would also seem to indicate that alumina -had no bad effect on a zinky slag. - - - - - ZINC OXIDE IN SLAGS - - BY W. MAYNARD HUTCHINGS - - (December 24, 1903) - - -From time to time, in various articles and letters on metallurgical -subjects in the _Engineering and Mining Journal_, the question of the -removal of zinc oxide in slags is referred to, and the question is -raised as to the form in which it is contained in the slags. - -I gather that opinion is divided as to whether zinc oxide enters into -the slags as a combined silicate, or whether it is simply carried into -them in a state of mechanical mixture. - -For many years I have taken great interest in the composition of slags, -and have studied them microscopically and chemically. The conclusion to -which I have been led as regards zinc oxide is, that in a not too basic -slag it is originally mainly, if not wholly, taken up as silicate along -with the other bases. On one occasion, one of my furnaces for several -days produced a slag in which beautiful crystals of willemite were -very abundant, both free in cavities and also imbedded throughout the -mass of solid slag, as shown in thin sections under the microscope. In -the same slag was a large amount of magnetite, all of which contained -a considerable proportion of zinc oxide combined with it. Magnetite -crystals, separated out from the slag and treated with strong acid, -yielded shells of material retaining the form of the original mineral, -rich in zinc oxide; an inter-crystallized zinc-iron spinel, in fact. -I have seen and separated zinc-iron spinels very rich in zinc oxide -from other slags. They have been seen in the slags at Freiberg; and -of course everybody knows the very interesting paper by Stelzner and -Schulze, in which they described the beautiful formations of spinels -and willemite in the walls of the retorts of zinc works. - -I think there is thus good ground for concluding that zinc oxide is -slagged off as combined silicate, and that free oxide does not exist -in slags; though zinc oxide does occur in them after solidification, -combined with other oxides, in forms ranging from a zinkiferous -magnetite to a more or less impure zinc-iron, or zinc-iron-alumina -spinel, these minerals having crystallized out in the earlier stages of -cooling. - -The microscope showed that the crystals of willemite, mentioned above, -were the first things to crystallize out from the molten slag. The main -constituent was well-crystallized iron-olivine-fayalite. - - - - - PART V - - LIME-ROASTING OF GALENA - - - - - THE HUNTINGTON-HEBERLEIN PROCESS - - (July 6, 1905) - - -It is a fact, not generally known, that the American Smelting and -Refining Company is now preparing to introduce the Huntington-Heberlein -process in all its plants, this action being the outcome of extensive -experimentation with the process. It is contemplated to employ the -process not only for the desulphurization of all classes of lead ore, -but also of mattes. This is a tardy recognition of the value of a -process which has been before the metallurgical profession for nine -years, the British patent having been issued under date of April -16, 1896, and has already attained important use in several foreign -countries; but it will be the grandest application in point of -magnitude. - -The Huntington-Heberlein is the first of a new series of processes -which effect the desulphurization of galena on an entirely new -principle and at great advantage over the old method of roasting. -They act at a comparatively low temperature, so that the loss of lead -and silver is reduced to insignificant proportion; they eliminate the -sulphur to a greater degree; and they deliver the ore in the form of -a cinder, which greatly increases the smelting speed of the blast -furnace. They constitute one of the most important advances in the -metallurgy of lead. The roasting process has been the one in which -least progress has been made, and it has remained a costly and wasteful -step in the treatment of sulphide ores. In reducing upward of 2,500,000 -tons of ore per annum, the American Smelting and Refining Company is -obliged to roast upward of 1,000,000 tons of ore and matte. - -The Huntington-Heberlein process was invented and first applied at -Pertusola, Italy. It has since been introduced in Germany, Spain, Great -Britain, Mexico, British Columbia, Tasmania, and Australia, in the last -at the Port Pirie works of the Broken Hill Proprietary Company. Efforts -were made to introduce it in the United States at least five years ago, -without success and with little encouragement. The only share in this -metallurgical improvement that this country can claim is that Thomas -Huntington, one of the inventors, is an American citizen, Ferdinand -Heberlein, the other, being a German. - - - - - LIME-ROASTING OF GALENA - - (September 22, 1905) - - -The article of Professor Borchers (see p. 116) is, we believe, the -first critical discussion of the reactions involved in the new methods -of desulphurizing galena, as exemplified in the processes of Huntington -and Heberlein, Savelsberg, and Carmichael and Bradford, although -the subject has been touched upon by Donald Clark, writing in the -_Engineering and Mining Journal_. It is perfectly obvious from a study -of the metallurgy of these processes that they introduce an entirely -new principle in the oxidation of galena, as Professor Borchers points -out. Inasmuch as there are already three of these processes and are -likely to be more, it will be necessary to have a type-name for this -new branch of lead metallurgy. We venture to suggest that it may be -referred to as the “lime-roasting of galena,” inasmuch as lime is -evidently a requisite in the process; or, at all events, it is the -agent which will be commonly employed. - -When the Huntington-Heberlein process was first described, it did not -even appear a simplification of the ordinary roasting process, but -rather a complication of it. The process attracted comparatively little -attention, and was indeed regarded somewhat with suspicion. This was -largely due to the policy of the company which acquired the patent -rights in refusing to publish the technical information concerning it -that the metallurgical profession expected and needed. The history of -this exploitation is another example of the disadvantage of secrecy -in such matters. The Huntington-Heberlein process has only become -thoroughly established as a new and valuable departure in metallurgy, a -departure which is indeed revolutionary, nine years after the date of -the original patent. In proprietary processes time is a particularly -valuable element, inasmuch as the life of a patent is limited. - -From the outset the explanation of Huntington and Heberlein as to the -reactions involved in their process was unsatisfactory. Professor -Borchers points out clearly that their conception of the formation of -calcium peroxide was erroneous, and indicates strongly the probability -that the active agent is calcium plumbate. It is very much to be -regretted that he did not go further with his experiments on this -subject, and it is to be hoped that they will be taken up by the -professors of metallurgy in other metallurgical schools. The formation -of calcium plumbate in the process was clearly forecasted, however, by -Carmichael and Bradford in their first patent specification; indeed, -they considered that the sintered product consisted largely of calcium -plumbate. - -Even yet, we have only a vague idea of the reactions that occur in -these processes. There is undoubtedly a formation of calcium sulphate, -as pointed out by Borchers and Savelsberg; but that compound is -eventually decomposed, since it is one of the advantages of the -lime-roasting that the sintered product is comparatively low in -sulphur. Is it true, however, that the calcium eventually becomes -silicate? If so, under what conditions is calcium silicate formed? The -temperature maintained throughout the process is low, considerably -lower than that required for the formation of any calcium silicate by -fusion. - -Moreover, it is not only galena which is decomposed by the new -method, but also blende, pyrite and copper sulphides. The process is -employed very successfully in the treatment of Broken Hill ore that is -rather high in zinc sulphide, and it is also to be employed for the -desulphurization of mattes. What are the reactions that affect the -desulphurization of the sulphides other than lead? - -There is a wide field for experimental metallurgy in connection with -these new processes. The important practical development is that they -do actually effect a great economy in the reduction of lead sulphide -ores. - - - - - THE NEW METHODS OF DESULPHURIZING GALENA[18] - - BY W. BORCHERS - - (September 2, 1905) - - -An important revolution in the methods of smelting lead ore, which had -to a large extent remained for centuries unchanged in their essentials, -was wrought by the invention of Huntington and Heberlein in 1896. More -especially is this true of the roast-reduction method of treating -galena, which consists of oxidizing roasting in a reverberatory furnace -and subsequent smelting of the roasted product in a shaft furnace. - -The first stage of the roast-reduction process, as carried out -according to the old method, viz., the oxidizing roast of the galena, -serves to convert the lead sulphide into lead oxide: - - PbS + 3O = PbO + SO₂. - -Owing to the basic character of the lead oxide, the production of a -considerable quantity of lead sulphate was of course unavoidable: - - PbO + SO₂ + O = PbSO₄. - -As this lead sulphate is converted back into sulphide in the -blast-furnace operation, and so adds to the formation of matte, it -has always been the aim (in working up ores containing little or no -copper to be concentrated in the matte) to eliminate the sulphate as -completely as possible, by bringing the charge, especially toward the -end of the roasting operation, into a zone of the furnace wherein -the temperature is sufficiently high to effect decomposition of the -sulphate by silica: - - PbSO₄ + SiO₂ = PbSiO₃ + SO₃. - -But in the usual mode of carrying out the roast in reverberatory -furnaces, the roasting itself on the one hand, and the decomposition of -the sulphates on the other, were effected only incompletely and with -widely varying results. - -Little attention has been paid in connection with the roast-reduction -process to the reaction between sulphates and undecomposed sulphides, -which plays so important a part in the roast-reaction method of lead -smelting. As is well known, lead sulphate reacts with lead sulphide in -varying quantities, forming either metallic lead or lead oxide, or a -mixture of both. A small quantity of lead sulphate reacting with lead -sulphide yields under certain conditions only lead: - - PbSO₄ + PbS = Pb₂ + 2SO₂. - -Within certain temperature limits this reaction even proceeds with -liberation of heat. In order to encourage it, it is necessary to create -favorable conditions for the formation of considerable quantities -of sulphate right at the beginning of the operation. This was first -achieved by Huntington and Heberlein, but not in the simplest nor in -the most efficient manner. And, indeed, the inventors were not by any -means on the right track as to the character of their process, so far -as the chemical reactions involved are concerned. - -At first sight the Huntington-Heberlein process does not even appear -as a simplification, but rather as a complication, of the roasting -operation. For in place of the roast carried out in one apparatus -and continuously, there are two roasts which have to be carried out -separately and in two different forms of apparatus; nevertheless, the -ultimate results were so favorable that the whole process is presumably -acknowledged, without reservation, by all smelters as one of the most -important advances in lead smelting. - -It is useful to examine in the light of the German patent specification -(No. 95,601 of Feb. 28, 1897) what were the ideas of its originators -regarding the operation of this process and the reactions leading to -such remarkable results. They stated: - -“We have made the observation that when powdered lead sulphide (PbS), -mixed with the powdered oxide of an alkaline earth metal, _e.g._, -calcium oxide, is exposed to the action of air at bright red heat -(about 700 deg. C.), and is then allowed to cool without interrupting -the supply of air, an oxidizing decomposition takes place when dark-red -heat (about 500 deg. C.) is reached, sulphurous acid being expelled, -and a considerable amount of heat evolved; if sufficient air is then -continuously passed through the charge, dense vapors of sulphurous acid -escape, and the mixture gradually sinters together to a mass, in which -the lead of the ore is present in the form of lead oxide, provided the -air blast is continued long enough; there is no need to supply heat in -this process—the heat liberated in the reaction is quite sufficient to -keep it up.” - -The inventors explained the process as follows: - -“At a bright-red heat the calcium oxide (CaO) takes up oxygen from -the air supplied, forming calcium peroxide (CaO₂), which latter -afterward, in consequence of cooling down to dark-red heat, again -decomposes into monoxide and oxygen; this nascent oxygen oxidizes a -part of the lead sulphide to lead sulphate, which then reacts with a -further quantity of lead sulphide, with evolution of sulphur dioxide -and formation of lead oxide.” - -Assuming the formation of calcium peroxide (CaO₂), the process -leading to the desulphurization would therefore be represented as -follows: - - 1. at 700° C. CaO + O = CaO₂ - 2. at 500° C. 4CaO₂ + PbS = 4CaO + PbSO₄ - 3. at the melting point PbS + PbSO₄ = 2PbO + 2SO₂ (?) - -Reactions 1 and 2 combined, assuming the presence of sufficient oxygen, -give: - - PbS + 4CaO + 4O = PbSO₄ + 4CaO. - -Now the invention consists in applying the observation described above -to the working up of galena, and other ores containing lead sulphide, -for metallic lead; and the essential novelty of the process therefore -consists in passing air through the mass cooled to a dark-red heat (500 -deg. C.). - -This feature sharply distinguishes it from other known processes. -It is true that in previous processes (compare the Tarnowitz -reverberatory-furnace process, the roasting process used at -Munsterbusch near Stolberg, and others) the lead ore was mixed with -limestone or dolomite (which are converted into oxides in the early -stage of the roast) and the heat was alternately raised and lowered; -but in all cases only a surface action of the air was produced, the air -supply being provided simply by the furnace draft. Passing air through -the mass cooled down, as indicated above, leads to the important -economic advantages of reducing the fuel consumption, the losses of -lead, the manual labor (raking) and the dimensions of the roasting -apparatus. - -In order to carry out the process of this invention, the powdered ore -is intimately mixed with a quantity of alkaline earth oxide, _e.g._, -calcium oxide, corresponding to its sulphur content; if the ore -already contains alkaline earth, the quantity to be added is reduced -in accordance. The mixture is heated to bright-red heat (700 deg. C.) -in the reverberatory furnace, in a strongly oxidizing atmosphere, is -then allowed to cool down to dark-red heat (500 deg. C.), also in -strongly oxidizing atmosphere, is transferred to a vessel called the -“converter,” and atmospheric air is passed through at a slight pressure -(the inventors have found a blast corresponding to 35 to 40 cm. head -of water suitable).[19] The heat liberated is quite sufficient to keep -the charge at the reaction temperature, but, if desired, hot blast may -also be used. The mixture sinters together, and (while sulphurous acid -gas escapes) it is gradually converted into a mass consisting of lead -oxide, gangue and calcium sulphate, from which the lead is extracted in -the metallic form, by any of the known methods, in the shaft furnace. -The operation is concluded as soon as the mass, by continued sintering, -has become impermeable to the blast. If the operation is properly -conducted, the gas escaping contains only small quantities of volatile -lead compounds, but on the other hand up to 8 per cent. by volume of -sulphur dioxide. This latter can be collected and further worked up. - -“In place of the oxide of an alkaline earth, ferrous oxide (FeO) or -manganous oxide (MnO) may also be used.” - -According to the reports on the practice of this process which have -been published,[20] conical converters of about 1700 mm. (5 ft. 6 -in.) upper diameter and 1500 mm. (5 ft.) depth are used in Australian -works. At a new plant at Port Pirie (Broken Hill Proprietary Company) -converters 2400 mm. (7 ft. 10 in.) in diameter and 1800 mm. (5 ft. -11 in.) deep have been installed. These latter will hold a charge of -about eight tons. In the lower part of these converters, at a distance -of about 600 mm. (2 ft.) from the bottom, there is placed an annular -perforated plate, and upon this a short perforated tube, closed above -by a plate having only a limited number of holes. - -No details have been published with regard to the European -installations. The general information which the Metallurgische -Gesellschaft[21] placed at my disposal upon request some years ago, -for use in my lecture courses, was restricted to data regarding the -consumption of fuel and labor in roasting and smelting the ores, which -was figured at about one-third or one-half of the consumption in the -former processes, to the demonstration of the large output of the -comparatively small converters, and to the reduced size of the roasting -plant as the result. But the European establishments which introduced -this process were bound by the owners of the patents, notwithstanding -the protection afforded by the patents, to give no information whatever -regarding the process to outsiders, and not to allow any inspection of -the works. - -On the other hand, a great deal appeared in technical literature -which was calculated to excite curiosity. Moreover, as professor of -metallurgy, it was my duty to instruct my pupils concerning this -process among others, and it was therefore very gratifying to me -that one of the students in my laboratory took a special interest in -the treatment of lead ore. I gave him opportunity to install a small -converter, in order to carry out the process on a small scale, and -in spite of the slender dimensions of the apparatus the very first -experiments gave a complete success. - -However, I could not harmonize the explanation of the process given by -the inventors with the knowledge which I had acquired in my many years’ -practical experience in the manufacture of peroxides. It is clear from -the patent specification that in the roasting operation at 700 deg. -C. a compound must be formed which functions as an excellent oxygen -carrier, for on cooling to 500 deg. C. the further oxidation then -proceeds to the end not only without any external application of heat, -but even with vigorous evolution of heat. No more striking instance -than this could be desired by the theorists who have of recent years -again become so enthusiastic over the idea of catalysis. Huntington -and Heberlein regarded calcium peroxide as the oxygen carrier, but that -is a compound which cannot exist at all under the conditions which -obtain in their process. The peroxides of the alkaline earths are so -very sensitive that in preparing them the small quantities of carbon -dioxide and water must be extracted carefully from the air, and yet -in the process, in an atmosphere pregnant with carbon dioxide, water, -sulphurous acid, etc., calcium peroxide, the most sensitive of the -whole group, is supposed to form! This could not be. - -The only compounds known as oxygen carriers, and capable of existing -under the conditions of the process, are calcium plumbate and plumbite. -I have emphasized this point from the first in my lectures on -metallurgy, when dealing with the Huntington-Heberlein process, and, in -point of fact, this assumption has since been proved to be correct by -the work of L. Huppertz, one of my students. - -During my practical activity (1879-1891) I had prepared barium peroxide -and lead peroxide in large quantities on a manufacturing scale, the -last-mentioned through the intermediate formation of plumbites and -plumbates: - - 2NaOH + PbO + O = Na₂PbO₃ + H₂O - -or: - - 4NaOH + PbO + O = Na₄PbO₄ + 2H₂O. - -An experiment made in this connection showed that calcium plumbate is -formed just as readily from slaked lime and litharge as the sodium -plumbates above. Litharge is an intermediate product, produced in -large quantities in lead works, and must in any case be brought -back into the process. If, then, the litharge is roasted at a low -temperature with slaked lime, the roasting of the galena could perhaps -be entirely avoided by introducing that ore together with calcium -plumbate into the converter, after the latter had once been heated up. -Mr. Huppertz undertook the further development of this process, but I -have no information on the later experimental results, as he placed -himself in communication with neighboring lead works for the purpose -of continuing his investigation, and has not since then given me any -precise data. I will therefore confine myself to the statement that -the fundamental idea for the experiments, which Mr. Huppertz undertook -at my suggestion, was the following: - -To dispense with the roasting of the galena, which is necessary -according to Huntington and Heberlein; in other words, to convert -the galena by direct blast, with the addition of calcium plumbate, -the latter being produced from the litharge which is an unavoidable -intermediate product in the metallurgy of lead and silver. (Borchers, -“Elektrometallurgie,” 3d edition, 1902-1903, p. 467.) - -This alone would, of course, have meant a considerable simplification -of the roast, but the problem of the roasting of galena has been solved -in a better way by A. Savelsberg, of Ramsbeck, Westphalia, who has -determined the conditions for directly converting the galena with the -addition of limestone and water and without previous roasting. He has -communicated the following information regarding these conditions: - -In order that, in blowing the air through the mixture of ore and -limestone, an alteration of the mixture may not take place owing to the -lighter particles of the limestone being carried away, it is necessary -(quite at variance with the processes in use hitherto, in which for the -sake of economy stress is laid on the precaution of charging the ore -as dry as possible into the apparatus) to add a considerable quantity -of water to the charge before introducing it into the converter. The -water serves this purpose perfectly, also preventing any change in -the mixture of ore and limestone, which invariably occurs if the ore -is used dry. The water, moreover, exerts a very beneficial action -in the process, inasmuch as it aids materially in the formation and -temporary retention of sulphuric acid, which latter then, by its -oxidizing action, greatly enhances the reaction and consequently the -desulphurization of the ore. Furthermore, the water tends to moderate -the temperature in the charge by absorbing heat in its volatilization. - -In carrying out the process the converter must not be filled entirely -all at once, but first only in part, additional layers being charged -in gradually in the course of the operation. In this way a uniform -progress of the reaction in the mass is secured. - -The following mode of procedure is advantageously adopted: A small -quantity of glowing fuel (coal, coke, etc.) is introduced into the -converter, which is provided at the bottom with a grate (perforated -sheet iron), the grate being first covered with a thin layer of crushed -limestone in order to protect it from the action of the red-hot coals -and ore. Upon this red-hot fuel a uniform layer of the wetted mixture -of crude ore and limestone is placed. When the surface of the first -layer has acquired a uniform red heat, a fresh layer is charged on, -and this is continued, layer by layer, until the converter is quite -full. While the layers are still being put on, the blast is passed in -at quite a low pressure, and only when the converter is entirely filled -is the whole force of the blast, at a rather greater pressure, turned -on. There then sets in a kind of slag formation, which, however, is -preceded by a very vigorous desulphurization. After the termination of -the process, which can be recognized by the fact that vapors cease to -be evolved, and that the surface of the ore becomes hard, the converter -is tipped over, and the desulphurized mass drops out as a solid cone of -slag, which is then suitably broken up for the subsequent smelting in -the shaft furnace. - -Savelsberg explains the reaction of this process as follows: - -“1. The particles of limestone act mechanically, gliding in between the -particles of lead ore and separating them from one another. In this -way a premature sintering is prevented, and the whole mass is rendered -loose and porous. - -“2. The limestone moderates the reaction temperature produced in the -combustion of the sulphur, so that the fusion of the galena, the -formation of dust and the separation of metallic lead are avoided, -or at least kept within the limits permissible. The lowering of the -temperature of reaction is due partly to the decomposition of the -limestone into caustic lime and carbon dioxide, in which heat is -absorbed, and partly to the consumption of the quantity of heat which -is necessary in the further progress of the operation for the formation -of a slag from the gangue of the ore and the lead oxide produced. - -“3. The limestone gives rise to chemical reactions. By its -decomposition it produces lime, which, at the moment of its formation, -is converted into calcium sulphate at the expense of the sulphur -in the ore. The calcium sulphate at the time of slag formation is -converted into silicate by the silica present, sulphuric acid being -evolved. The limestone therefore assists directly and forcibly in the -desulphurization of the ore, causing the formation of sulphuric acid at -the expense of the sulphur in the ore, the sulphuric acid then acting -as a strong oxidizing agent toward the sulphur in the ore.” - -The most conclusive proof for the correctness of the opinion which I -expressed above, that it is very important to create at the beginning -of the operation the conditions for the formation of as much sulphate -as possible, has been furnished by Carmichael and Bradford. They -recommend that gypsum be added to the charge in place of limestone. At -one of the works of the Broken Hill Proprietary Company (where their -process has been carried on successfully, and where lead ores very rich -in zinc had to be worked up) the dehydrated gypsum was mixed with an -equal quantity of concentrate and three times the quantity of slime -from the lead ore-dressing plant, as in the table given herewith: - - ─────────────────┬────────┬─────────────┬──────────┬──────── - │ OF THE │ OF THE │ OF THE │ OF THE - CONTENTS │ SLIME │ CONCENTRATE │ CALCIUM │ WHOLE - │ │ │ SULPHATE │ CHARGE - ─────────────────┼────────┼─────────────┼──────────┼──────── - Galena │ 24 │ 70 │ │ 29 - Zinc blende │ 30 │ 15 │ │ 21 - Pyrites │ 3 │ │ │ 2 - Ferric oxide │ 4 │ │ │ 2.5 - Ferrous oxide │ 1 │ │ │ 1 - Manganous oxide │ 6.5 │ │ │ 5 - Alumina │ 5.5 │ │ │ 3 - Lime │ 3.5 │ │ 4.1 │ 10 - Silica │ 23 │ │ │ 14 - Sulphur trioxide │ │ │ 59 │ 12 - ─────────────────┴────────┴─────────────┴──────────┴──────── - -The charge is mixed, with addition of water, in a suitable pug-mill. -The mass is then, while still wet, broken up into pieces 50 mm. (2 in.) -in diameter, which are then allowed to dry on a floor in contact with -air; in doing so they set hard, owing to the rehydration of the gypsum. - -As in the case of the Savelsberg process, the converters are heated -with a small quantity of coal, are filled with the material prepared -in the manner above described, and the charge is blown, regulating -the blast in such manner that, after the moisture present has been -dissipated, a gas of about 10 per cent. SO₂ content is produced, -which is worked up for sulphuric acid in a system of lead chambers. - -The reactions are in this case the same as in the Savelsberg process, -for here also calcium sulphate is formed transitorily, which, like -other sulphates, reacts partly with sulphides, partly with silica. - -Where gypsum is available and cheap, the Carmichael-Bradford process -must be given preference; in all other cases unquestionably the -Savelsberg process is superior, owing to its great simplicity. - - - - - LIME-ROASTING OF GALENA - - BY W. MAYNARD HUTCHINGS - - (_October 21, 1905_) - - -Much interest attaches to the paper by Professor Borchers, recently -presented in the _Engineering and Mining Journal_ (Sept. 2, 1905) on -“New Methods of Desulphurizing Galena,” together with an editorial on -“Lime-Roasting of Galena”; it is a curious coincidence that the same -issue contained also an article on the “Newer Treatment of Broken Hill -Sulphides,” in which is shown the importance of the new methods as a -contribution to actual practice. - -For some years it had been a source of surprise to me that a new -process, so interesting and so successful as the Huntington-Heberlein -treatment of sulphide ores, should have received scarcely any notice -or discussion. This lack, however, now appears to be remedied. The -suggestion that the subject should be discussed in the _Journal_ -is good, as is also that of the designation “Lime-Roasting” for a -type-name. Such observations and experiments on the subject as I have -had occasion to record have, for many years, figured in my note-books -under that heading. - -Whatever may be the final results of the later processes, now before -the metallurgical world or still to come, there can be no doubt -whatever that full and exclusive credit must be given to Huntington and -Heberlein, not only for first drawing attention to the use of lime, but -also for working out and introducing practically the process. It has -been a success from the first; and so far as part of it is concerned, -it seems to be an absolute and fundamental necessity which later -inventors can neither better nor set aside. The other processes, since -patented, however good they may be, are simply grafts on this parent -stem. - -It is, however, quite certain that Huntington and Heberlein, in the -theoretical explanation of the process, failed to understand the most -important reactions. Their attributing the effect to the formation and -action of calcium peroxide affords a sad case of _a priori_ assumption -devoid of any shred of evidence. As Professor Borchers points out, -calcium peroxide, so difficult to produce and so unstable when formed, -is an absolute and absurd impossibility under the conditions in -question. Probably many rubbed their eyes with astonishment on reading -that part of the patent on its first appearance, and hastened to look -up the chemical authorities to refresh their minds, lest something as -to the nature of calcium peroxide might have escaped them. - -Fortunately the patent law is such that there was no danger of a really -good and sound invention being invalidated by a wrong theoretical -explanation by its originators. But, nevertheless, it was a misfortune -that the inventors did not understand their own process. Had they -known, they could have added a few more words to their patent-claims -and rendered the Carmichael patent an impossibility. - -Professor Borchers appears to consider that the active agent in the new -process is calcium plumbate. That this compound may play a part at some -stage of the process may be true; this long ago suggested itself to -some others. We may yet expect to hear that the experiments undertaken -by Professor Borchers himself, and by others at his instigation (in -which calcium plumbate is separately prepared and then brought into -action with lead sulphide), have given good results. But it does not -appear so far that there is any real proof that calcium plumbate is -formed in the Huntington-Heberlein or other similar processes; and it -is difficult to see at what stage or how it would be produced under the -conditions in question. This is a point which research may clear up, -but it should not be taken for granted at this stage. Indeed, it seems -to me that the results obtained may be fairly well explained without -calling calcium plumbate into play at all. - -Of course the action of lime in contact with lead sulphide excited -interest many years before the new process came into existence. My own -attention to it dates back more than a dozen years before that time (I -was in charge of works where I found the old “Flintshire process” still -in use). - -Percy pointed out, in his work on lead smelting, that on the addition -of slaked lime to the charge, at certain stages, to “stiffen it up,” -the mixture could be seen to “glow” for a time. When I myself saw this -phenomenon, I commenced to make some observations and experiments. -Also (as others probably had done), I had observed that charges of lead -with calcareous gangue are roasted more rapidly and better than others, -and to an extent which could not be wholly explained by simple physical -action of the lime present. - -Simple experiments made in assay-scorifiers in a muffle, on lime -roasting, are very striking, and I think quite explain a good part -of what takes place up to a certain stage in the processes now under -consideration. I tried them a number of years ago, on many sorts of -ore, and again more recently, when studying the working of the new -patents. For illustration, I will take one class of ore (Broken Hill -concentrate), using a sample assaying; Pb, 58 per cent.; Fe, 3.6 per -cent.; S, 14.6 per cent.; SiO₂, 3 per cent. The ore contained some -pyrite. If two scorifiers are charged, one with the finely powdered -ore alone, and one with the ore intimately mixed with, say, 10 per -cent. of pure lime, and placed side by side just within a muffle at -low redness, the limed charge will soon be seen to “glow.” Before the -simple ore charge shows any sign of action, the limed charge rapidly -ignites all over, like so much tinder, and heats up considerably above -the surrounding temperature, at the same time increasing noticeably -in bulk. This lasts for some time, during which hardly any SO₂ -passes off. After the violent glowing is over, the charge continues -to calcine quietly, giving off SO₂, but is still far more active -than its neighbor. If, finally, the fully roasted charge is taken out, -cooled and rubbed down, it proves to contain no free lime at all, but -large quantities of calcium sulphate can be dissolved out by boiling in -distilled water. For instance, in one example where weighed quantities -were taken of lime and the ore mentioned, the final roasted material -was shown to contain nearly 23 per cent. of CaSO₄; the quantity -actually extracted by water was 20.2 per cent. Further tests show -that the insoluble portion still contains calcium sulphate intimately -combined with lead sulphate, but not extractable by water. - -There is no doubt that when lead sulphide (or other sulphide) is -heated with lime, with free access of air, the lime is rapidly and -completely converted into sulphate. The strong base, lime, apparently -plays the part of “catalyzer” in the most vigorous manner, the first -SO₂ evolved being instantly oxidized and combined with the lime -to sulphate, with so strong an evolution of heat that the operation -spreads rapidly and still goes on energetically, even if the scorifier -is taken out of the muffle. Also, the “catalytic” action starts the -oxidation of the sulphides at a far lower temperature than is required -when they are roasted alone. - -If, in place of lime, we take an equivalent weight of pure calcium -carbonate and intimately mix it with ore, we obtain just the same -action, only it takes a little longer to start it. Once started, it -is almost as vigorous and rapid, and with the same results. It does -not seem correct to assume (as is usually done) that the carbonate has -first to be decomposed by heat, the lime then coming into action. The -reaction commences in so short a time and while the charge is still -so cool, that no appreciable driving off of CO₂ by heat only can -have taken place. The main liberation of the CO₂ occurs during the -vigorous exothermic oxidation of the mixture, and is coincident with -the conversion of the CaO into CaSO₄. - -If, in place of lime or its carbonate, we use a corresponding quantity -of pure calcium sulphate and mix it with the ore, we see very energetic -roasting in this case also, with copious evolution of sulphur dioxide, -only it is much more energetic and rapid and occurs at a lower -temperature than in the case of a companion charge of ore alone. - -It is very easily demonstrated that the CaSO₄ in contact with the -still unoxidized ore (whether it has been introduced ready made or has -been formed from lime or limestone added) greatly assists the further -roasting, in acting as a “carrier” and enabling calcination to take -place more rapidly and easily and at a lower temperature than would -otherwise be the case. - -The result of these experiments (whether we mix the ore with CaO, -CaCO₃, or CaSO₄) is that we arrive with great ease and rapidity -at a nearly dead-sweet roast. The lime is converted into sulphate, and -the lead partly to sulphate and partly to oxide. Two examples out of -several, both from the above ore, gave results as follows: - -No. 1—Roasted with 20 per cent. CaCO₃ (= 11.2 per cent. CaO); -sulphide sulphur, 0.02 per cent.; sulphate sulphur, 9.30 per cent.; -total sulphur, 9.32 per cent. - -No. 8—Roasted with 27.2 per cent. CaSO₄ (= 11 per cent. CaO); -sulphide sulphur, 0.05 per cent.; sulphate sulphur, 11.28 per cent.; -total sulphur, 11.33 per cent. - -If these calcined products are now intimately mixed with additional -silica (in about the proportions used in the Huntington-Heberlein -process) and strongly heated, fritting is brought about and the sulphur -content is reduced by the decomposition of the sulphates by the silica. -Thus, the resultant material of experiment No. 1, above, when treated -in this manner with strong heat for three hours, was sintered to a mass -which was quite hard and stony when cold, and which contained 6.75 -per cent. of total sulphur. Longer heating drives out more sulphur, -but a very long time is required; in furnaces, and on a large scale, -it is with great difficulty and cost that a product can be obtained -comparable with that which is rapidly and cheaply turned out from the -“converters” of the new process. - -To return to the Huntington-Heberlein process, working, for example, -on an ore more or less like the one given above, we may assume that, -during the comparatively short preliminary roast, the lime is all -rapidly converted into CaSO₄ and that some PbSO₄ is also formed -(but not much, as the mixture to be transferred from the furnace to -the converter requires not less than 6 to 8 per cent. of sulphur to -be still present as sulphide, in order that the following operation -may work at its best). As the blast permeates the mass, oxidation is -energetic; no doubt that CaSO₄ here plays a very important part -as a carrier of oxygen, in the same manner as we can see it act on a -scorifier or on the hearth of a furnace. - -What the later reactions are does not seem so clear. They are quite -different from those on the scorifier or on the open hearth of a -furnace, and result in the rapid formation (in successive layers of -the mixture, from the bottom upward) of large amounts of lead oxide, -fluxing the silica and other constituents to a more or less slaggy -mass, which decomposes the sulphates and takes up the CaO into a -complex and easily fused silicate. It is true that, as a whole, the -contents of a well-worked converter are never very hot, but locally -(in the regions where the progressive reaction and decomposition from -below upward is going on) the temperature reached is considerable. This -formation of lead oxide is so pronounced at times that one may see in -the final product considerable quantities of pure uncombined litharge. - -When the work is successful, the mass discharged from the converters -is a basic silicate of PbO, CaO, and oxides of other metals present, -and nearly all the sulphates have disappeared. A large piece of yellow -product (which was taken from a well-worked converter) contained only -1.1 per cent. of total sulphur. - -It may be that calcium plumbate is formed and plays a part in these -reactions; but its presence would be difficult to prove, and its -formation and existence during these stages would not be easy to -explain. Neither does it seem necessary, as the whole thing appears to -be capable of explanation without it. - -While the mixture in the converter is still dry and loose, energetic -oxidation of the sulphides goes on, with the intervention of the -CaSO₄ as a carrier. As soon as the heat rises sufficiently, fluxing -commences in a given layer and sulphates are decomposed. The liberated -sulphuric anhydride, at the locally high temperature and under the -existing conditions, will act with the greatest possible vigor on the -sulphides in the adjacent layers; these layers will then in their -turn flux and act on those above them, till the whole charge is -worked out. The column of ore is of considerable hight, requiring a -blast of 1½ lb., or perhaps more, in the larger converters now used. -This pressure of the oxidizing blast (and of the far more powerfully -oxidizing sulphuric anhydride, continuously being liberated within the -mass of ore, locally very hot) constitutes a totally different set of -conditions from those obtained on the hearth of a furnace with the ore -in thin layers, where it is neither so hot nor under any pressure. -It is to these conditions, in which we have the continued intense -action of red-hot sulphuric anhydride under a considerable pressure -(together with the earlier action of the CaSO₄), that the remarkable -efficiency of the process seems to me to be due. - -In the Carmichael process, the preliminary roast is done away with, -CaSO₄ being added directly instead of having to be formed during the -operation from CaO and the oxidized sulphur of the ore. The charge in -the converter has to be started by heat supplied to it, and the work -then goes forward on the same lines as in the Huntington-Heberlein -process, so that we may assume that the reactions are the same and come -under the same explanation. - -Carmichael was quick to see what was really an important part and a -correct explanation of the original process. He was not misled by wrong -theory about any mythical calcium peroxide, and so he obtained his -patent for the use of CaSO₄ and the dispensing of the roast in a -furnace. - -This process would always be limited in its application by the -comparative rarity of cheap supplies of gypsum, but it appears to be -a great success at Broken Hill; there it is not only of importance in -working the leady ores, but also for making sulphuric acid for the new -treatment of mixed sulphides by the Delprat and Potter methods. For -this purpose, the use of CaSO₄ will have the additional advantage -that the mixture to be worked in the converter will contain not -only the sulphur of the ore, but also that of the added gypsum; on -decomposition, it will yield stronger gases for the lead chambers of -the acid plant. - -Finally comes the Savelsberg patent, which is the simplest of all; -not only (like the Carmichael process) avoiding the preliminary roast -with its extra plant, but also not requiring the use of ready-made -CaSO₄, as it uses raw ore and limestone directly in the converter. -I have no knowledge as to actual results of this process; and, so -far as I am aware, nothing on the subject has been published. But -Professor Borchers evidently has some information about it, and -regards it as the most successful of the methods of carrying out the -new ideas. On the face of it, there seems no reason why it should not -attain all the results desired, as the chemical and physical actions -of the CaO, and of the CaSO₄ formed from it, should come into play -in the same manner and in the same order as in the original process; -as it is carried out in the identical converter used by Huntington -and Heberlein, the final reactions (as suggested above) will take -place under the same conditions as to continuous decomposition _under -considerable heat and pressure_, which I regard as the most vital part -of the whole matter. - -It is well to emphasize again the fact that the idea, and the means of -obtaining these vital conditions, owe their origination to Huntington -and Heberlein. - - - - - THEORETICAL ASPECTS OF LEAD-ORE ROASTING[22] - - BY C. GUILLEMAIN - - (March 10, 1906) - - -It is well known that the process of roasting lead ores in -reverberatory furnaces proceeds in various ways according to the -composition of the ore in question. Thus in roasting a sulphide lead -ore rich in silica, one of the reactions is: - - PbS + 3O = PbO + SO₂. - -But this reaction is incomplete, for the gases which pass on in the -furnace are rich in SO₂ and in SO₃. And so it is found that -whatever lead oxide is formed passes over almost immediately into lead -sulphate, according to the reaction: - - PbO + SO₂ + O = PbSO₄. - -This reaction is the chief one which takes place. Whether the silicious -gangue serves as a catalyzer for the sulphur dioxide, or whether it -serves merely to keep the galena open to the action of the gases, the -end result of the roast is usually the formation of lead sulphate -according to the above reaction. - -In the case of an ore rich in galena, a slow roast is essential, for it -is desired to have the following reaction take place during the latter -part of the roast: - - PbS + 3PbSO₄ = 4PbO + 4SO₂. - -Now, if the heating were too rapid, not enough lead sulphate would be -found to react with the unaltered galena. The quick roasting of a rich -ore would result in the early sintering of the charge, and sintering -prevents the further formation of lead sulphate. Whether this sintering -(which takes place so easily and which is so harmful in the latter part -of the process) is due to the low melting point of the lead sulphide, -whether the heat evolved by the reaction - - PbS + 3O = PbO + SO₂ - -is sufficient to melt the lead sulphide, or whether other -thermochemical effects (notably the preliminary sulphatizing of the -lead sulphide) come into play, must for the present be undecided. -Suffice it to say that the sintering of the charge works against a good -roast. - -In the Tarnowitz process a definite amount of lead sulphide is -converted into lead sulphate by a preliminary roast. The sulphate then -reacts with the unaltered lead sulphide, and metallic lead is set free, -thus: - - PbS + PbSO₄ = 2Pb + 2SO₂. - -But when a very little of the sulphide has been transformed into -sulphate, and when there is so little of the latter present that only -a small amount of lead sulphide can be reduced to metallic lead, the -mass of ore begins to sinter and grow pasty. Very little lead could be -formed were it not for the addition of crushed lime to the charge just -before the sintering begins. This lime breaks up the charge and cools -it, prevents any sintering, and allows the continued formation of lead -sulphate. - -It scarcely can be held that the lime has any chemical effect in -forming lead sulphate, or in forming a hypothetical compound of lead -and calcium. Even if such theories were tenable from a physico-chemical -point of view, they would be lessened in importance by the fact that -other substances, such as purple ore or puddle cinder, act just as well -as the lime. - -There are now to be mentioned several new processes of lead-ore -roasting whose operations fall so far outside the common ideas on -the subject that their investigation is full of interest. For a long -time the attempt had been made to produce lead directly by blowing -air through lead sulphide in a manner analogous to the production of -bessemer steel or the converting of copper matte. In the case of the -lead sulphide, the oxidation of the sulphur was to furnish the heat -necessary to carry on the process. - -After many attempts along this line, Antonin Germot has perfected a -method wherein, by blowing air through molten galena, metallic lead -is obtained.[23] About 60 per cent. of a previously melted charge of -galena is sublimed as lead sulphide, and the rest remains behind as -metallic lead. The disadvantages of the process are the difficulties of -collecting all of the sublimate and of working it up. Moreover, it is -impossible as yet to secure two products of which one is silver-free -and the other silver-bearing. The silver values are in both the -metallic lead and in the sublimed lead sulphide. - -While the process just described answers for pure galena, it fails -with ores which contain about 10 per cent. of gangue. In the case of -such ores, they form a non-homogeneous mass when melted, and the blast -penetrates the charge with difficulty. If the pressure is increased -the air forces itself out through tubes and canals which it makes for -itself, and the charge freezes around these passages. - -Messrs. Huntington and Heberlein have gone a little farther. Although -they are unable to obtain metallic lead directly, they prepare the ore -satisfactorily for smelting in the blast furnace, after their roasting -is completed. The inventors found that if lead sulphide is mixed with -crushed lime, heated with access of air, and then charged into a -converter and blown, the sulphur is completely removed in the form of -sulphur dioxide. The charge, being divided by the lime, remains open -uniformly to the passage of air, and sinters only when the sulphur is -eliminated. - -The inventors announce, as the theory of their process, that at 700 -deg. C. the lime forms a dioxide of calcium (CaO₂) which at 500 deg. -C. breaks down into lime (CaO) and nascent oxygen. This nascent oxygen -oxidizes the lead sulphide to lead sulphate according to the reaction: - - PbS + 4O = PbSO₄. - -Furthermore it is claimed that the heat evolved by this last reaction -is large enough to start and keep in operation a second reaction, namely - - PbS + PbSO₄ = 2PbO + 2SO₂. - -The theory, as just mentioned, cannot be accepted, and some of the -reasons leading to its rejection will be given. - -It is well established that the simple heating of lime with access of -air will not result in further oxidation of the calcium. The dioxide -of calcium cannot be formed even by heating lime to incandescence in -an atmosphere of oxygen, nor by fusing lime with potassium chlorate. -Moreover, calcium stands very near barium in the periodic system. And -as the dioxide of barium is formed at a low temperature and breaks -up on continued heating, it seems absurd to suppose that the dioxide -of calcium would act in exactly the opposite manner. Moreover, a -consideration of the thermo-chemical effects will disclose more -inconsistencies in the ideas of the inventors. The breaking up of -CaO₂ into CaO and O is accompanied by the evolution of 12 cal. The -reaction of the oxygen (thus supposed to be liberated) upon the lead -sulphide is strongly exothermic, giving up 195.4 cal. So much heat is -produced by these two reactions that, if the ideas of the inventors -were true, the further breaking up of the calcium dioxide would stop, -as the whole charge would be above 500 deg. C. It appears, then, that -the explanations suggested by Messrs. Huntington and Heberlein are -untrue. - -In the usual roasting process, as carried out in reverberatory -furnaces, it is well established that the gangue, and whatever -other substances are added to the ore, prevent mechanical locking -up of charge particles, since they stop sintering. It is not at all -improbable that in the new roasting process the chief, if not the only, -part played by the lime is the same as that played by the gangue in -reverberatory-furnace roasting. A few observations leading to this -belief will be given. - -It is known that other substances will answer just as well as lime -in this new roasting process. Such substances are manganese and iron -oxides. Not only these two substances, but in fact any substance which -answers the purpose of diminishing the local strong evolution of heat, -due to the reaction: - - PbS + 3O = PbO + SO₂, - -serves just as well as the lime. This fact is proved by exhaustive -experiments in which mixtures of lead sulphide on the one hand, and -quartz, crushed lead slags, iron slags, crushed iron ores, crushed -copper slags, etc., on the other hand, were used for blowing. All -these substances are such that any chemical action, analogous to the -splitting up of CaO₂, or the formation of plumbates as suggested -by Dr. Borchers, cannot be imagined. The time is not yet ripe, without -more experiments on the subject, to assert conclusively that there -is no acceleration of the process due to the formation of plumbates -through the agency of lime. But the facts thus far secured point out -that such reactions are, at least, not of much importance. - -Theoretical considerations point out that it ought to be possible to -avoid the injurious local increase of temperature during the progress -of this new roasting process, without having to add any substance -whatever. To explain: The first reaction taking place in the roasting is - - PbS + 3O = PbO + SO₂ + 99.8 cal. - -Now the heat thus liberated may be successfully dispersed if there is, -in simultaneous progress, the endothermic reaction: - - PbS + 3PbSO₄ = 4PbO + 4SO₂ - 187 cal. - -Hence if there could be obtained a mixture of lead sulphide and of -lead sulphate in the proportions demanded by the above reaction, then -such a mixture ought to be blown successfully to lead oxide without -the addition of any other substance. Such a process has, in fact, been -carried out. The original galena is heated until the required amount of -lead sulphate has been formed. Then the mixture of lead sulphide and -of lead sulphate is transferred to a converter and blown successfully -without the addition of any other substance. - -The adaptability of an ore to the process just mentioned depends on -the cost of the preliminary roast and the thoroughness with which it -must be done. As is known, when lead sulphide is heated with access of -air, it is very easy to form sintered incrustations of lead sulphate. -If these incrustations are not broken up, or if their formation is not -prevented by diligent rabbling, the further access of air to the mass -is prevented and the oxidation of the charge stops. If ores with such -incrustations are placed in the converter without being crushed, they -remain unaltered by the blowing. If the incrustations are too numerous -the converting becomes a failure. - -It has been found that the adoption of mechanical roasting furnaces -prevents this. Such furnaces appear to stop the frequent failures of -the blowing which are due to the lack of care on the part of the -workmen during the preliminary roasting. Moreover, in such mechanical -furnaces a more intimate mixture of the sulphide with the sulphate -is obtained, and the degree of the sulphatizing roast is more easily -controlled. - -As a summary of the facts connected with this new blowing process, it -may be stated that the best method of working can be determined upon -and adopted if one has in mind the fact that the amount of substance -(lime) to be added is dependent on: 1, the amount of sulphur present; -2, the forms of oxidation of this sulphur; 3, the amount of gangue -in the ore; 4, the specific heats of the gangue and of the substance -added; 5, the degree of the preparatory roasting and heating. - -For example, with concentrates which run high in sulphur, there is -required either a large amount of additional material, or a long -preliminary roast. The specific heat of the added material must be -high, and the heat evolved by the oxidation of the sulphur in the -preliminary roast must be dispersed. Oftentimes it is necessary to cool -the charge partially with water before blowing. On the other hand, if -the ore runs low in sulphur, the preliminary roast must be short, and -the temperature necessary for starting the blowing reactions must be -secured by heating the charge out of contact with air. Not only must no -flux be added, but oftentimes some other sulphides must be supplied in -order that the blowing may be carried out at all. - -The opportunity for the acquisition of more knowledge on this subject -is very great. It lies in the direction of seeing whether or not the -strong local evolution of heat cannot be reduced by blowing with gases -poor in oxygen rather than with air. Mixtures of filtered flue gases -and of air can be made in almost any proportion, and such mixtures -would have a marked effect upon the possibility of regulating the -progress of the oxidation of the various ores and ore-mixtures which -are met with in practice. - - - - - METALLURGICAL BEHAVIOR OF LEAD SULPHIDE AND CALCIUM SULPHATE[24] - - BY F. O. DOELTZ - - (January 27, 1906) - - -In his British patent,[25] for desulphurizing sulphide ores, A. D. -Carmichael states that a mixture of lead sulphide and calcium sulphate -reacts “at dull red heat, say about 400 deg. C.,” forming lead sulphate -and calcium sulphide, according to the equation: - - PbS + CaSO₄ = PbSO₄ + CaS. - -Judging from thermo-chemical data, this reaction does not seem -probable. According to Roberts-Austen,[26] the heats of formation (in -kilogram-calories) of the different compounds in this equation are as -follows: PbS = 17.8; CaSO₄ = 318.4; PbSO₄ = 216.2; CaS = 92. -Hence we have the algebraic sum: - - -17.8 - 318.4 + 216.2 + 92 = -28.0 cal. - -As the law of maximum work does not hold, experiment only can -decide whether this decomposition takes place or not. The following -experiments were made: - -_Experiment 1._—Coarsely crystalline and specially pure galena was -ground to powder. Some gypsum was powdered, and then calcined. The -powdered galena and calcined gypsum were mixed in molecular proportions -(PbS + CaSO₄), and heated for 1½ hours to 400 deg. C., in a stream -of carbon dioxide in a platinum resistance furnace. The temperature was -measured with a Le Chatelier pyrometer. The material was allowed to -cool in a current of carbon dioxide. - -The mixture showed no signs of reaction. Under the magnifying glass the -bright cube-faces of galena could be clearly distinguished. If any -reaction had taken place, in accordance with the equation given above, -no bright faces of galena would have remained. - -_Experiment 2._—A similar mixture was slowly heated, also in the -electric furnace, to 850 deg. C., in a stream of carbon dioxide, and -was kept at this temperature for one hour. - -It was observed that some galena sublimed without decomposition, being -redeposited at the colder end of the porcelain boat (7 cm. long), in -the form of small shining crystals. The residue was a mixture of dark -particles of galena and white particles of gypsum, in which no evidence -of any reaction was visible under the microscope. That galena sublimes -markedly below its melting point has already been noted by Lodin.[27] - -_Experiment 3._—In order to determine whether the inverse reaction -takes place, for which the heat of reaction is + 28.0 cal., the -following equations are given: - - PbSO₄ + CaS = PbS + CaSO₄; - - 216.2 - 92 + 17.8 + 318.4 = 28. - -A mixture of lead sulphate and calcium sulphide was heated in a -porcelain crucible in a benzine-bunsen flame (Barthel burner). The -materials were supplied expressly “for scientific investigation” by the -firm, C. A. F. Kahlbaum. - -The white mixture turned dark and presently assumed the color which -would correspond to its conversion into lead sulphide and calcium -sulphate. This experiment is easy to perform. - -_Experiment 4._—The same materials, lead sulphate and calcium sulphide, -were mixed in molecular ratio (PbSO₄ + CaS), and were heated for 30 -minutes to 400 deg. C., on a porcelain boat in the electric furnace, -in a current of carbon dioxide. The mixture was allowed to cool in a -stream of carbon dioxide, and was withdrawn from the furnace the next -day (the experiment having been made in the evening). - -The mixture showed a dark coloration, similar to that of the last -experiment; but a few white particles were still recognizable. The -material in the boat smelled of hydrogen sulphide. - -_Experiment 5._—A mixture of pure galena and calcined gypsum, in -molecular ratio (PbS + CaSO₄), was placed on a covered scorifier -and introduced into the hot muffle of a petroleum furnace, at 700 to -800 deg. C. The temperature was then raised to 1100 deg. C. - -From 5 g. of the mixture a dark-gray porous cake weighing 3.7g. was -thus obtained. There was some undecomposed gypsum present, recognizable -under the magnifying glass. No metallic lead had separated out. When -hot hydrochloric acid was poured over the mixture, it evolved hydrogen -sulphide. The fracture of the cake showed isolated shining spots. The -supposition that it was melted or sublimed galena was confirmed by -the aspect of the cake when cut with a knife; the surface showed the -typical appearance of the cut surface of melted galena. On cutting, the -cake was found to be brittle, with a tendency to crumble. On boiling -with acetic acid, a little lead went into solution. Wetting with water -did not change the color of the crushed cake. - -_Experiment 6._—In his experiments for determining the melting point -of galena, Lodin[28] found that, in addition to its sublimation at a -comparatively low temperature, the galena also undergoes oxidation if -carbon dioxide is used as the “neutral” atmosphere. Lodin was therefore -compelled to use a stream of nitrogen in his determination of the -melting point of galena. Now the temperature of experiment 2 (850 deg. -C.), described heretofore, is not as high as the melting point of -galena (which lies between 930 and 940 deg. C.); therefore experiment 2 -was repeated in a stream of nitrogen, so as to insure a really neutral -atmosphere. A mixture of galena and calcined gypsum in molecular -ratio (PbS + CaSO₄) was heated to 850 deg. C., was kept at this -temperature for one hour, and allowed to cool, the entire operation -being carried out in a stream of nitrogen. - -Again, galena had sublimed away from the hotter end of the porcelain -boat (6.5 cm. long), and had been partially deposited in the form of -small crystals of lead sulphide at the colder end. The material in -the boat consisted of a mixture of particles having the dark color -of galena, and others with the white color of gypsum, the original -crystals of gypsum and the bright surfaces of the lead sulphide being -distinctly recognizable under the magnifying glass. The loss in weight -was 1.9 per cent. - -_Experiment 7._—For the same reason as in 2, experiment 5 was also -repeated, using a current of nitrogen. A mixture of galena and -calcined gypsum, in molecular ratio (PbS + CaSO₄) was heated in a -porcelain boat to 1030 deg. C., in a platinum-resistance furnace, and -allowed to cool, being surrounded by a stream of nitrogen during the -whole period. - -Some sublimation of lead sulphide again took place. The mixture was -seen to consist of white particles of gypsum, and others dark, like -galena. The loss in weight was 3.5 per cent. The mixture had sintered -together slightly; with hot hydrochloric acid, it evolved hydrogen -sulphide. On boiling with acetic acid, a little lead (only a trace) -went into solution. There was, therefore, practically no lead oxide -present; no metallic lead had separated out. - -_Experiment 8._—In experiment 3, lead sulphate and calcium sulphide -were mixed roughly and by hand (i.e., not weighed out in molecular -ratio); in this experiment such a mixture of lead sulphate and calcium -sulphide in molecular ratio (PbSO₄ + CaS) was heated in a porcelain -crucible in a benzine-bunsen flame. It presently turned dark, and a -dark gray product was obtained, as in the former experiment. - -_Experiment 9._—In a mixture of lead sulphate and sodium sulphide in -molecular ratio (PbSO₄ + Na₂S), the constituents react directly -on rubbing together in a porcelain mortar. The mass turns dark gray, -with formation of lead sulphide and sodium sulphate. - -If a similar mixture is heated, it also turns dark gray. On lixiviation -with water, a solution is obtained which gives a dense white -precipitate with barium chloride. - -_Experiment 10._—If lead sulphate and calcium sulphide are rubbed -together in a mortar, the mass turns a grayish-black. - -_Conclusion._—From these experiments I infer that the reaction - - PbS + CaSO₄ = PbSO₄ + CaS - -does not take place, but, on the contrary, that when lead sulphate and -calcium sulphide are brought together, the tendency is to form lead -sulphide and calcium sulphate. - -Nevertheless, on heating a mixture of galena and gypsum in contact with -air, lead sulphate will be formed along with lead oxide; not, however, -owing to any double decomposition of the galena with the gypsum, but -rather to the formation of lead sulphate from lead oxide and sulphuric -acid produced by catalysis, thus: - - PbO + SO₂ + O = PbSO₄. - -This is the well-known process which always takes place in roasting -galena, the explanation of which was familiar to Carl Friedrich -Plattner. That the presence of gypsum has any chemical influence on -this process seems to be out of the question according to the above -experiments. - - - - - THE HUNTINGTON-HEBERLEIN PROCESS - - BY DONALD CLARK - - (October 20, 1904) - - -The process was patented in 1897, and is based on the fact that galena -can be desulphurized by mixing it with lime and blowing a current of -air through the mixture. If the temperature is dull red at the start, -no additional source of heat is necessary, because the reaction causes -a great rise in temperature. The chemistry of the process cannot be -said at present to have been worked out in detail. - -The reactions given by the patentees are not satisfactory, since -calcium dioxide is formed only at low temperatures and is readily -decomposed on gently warming it; lead oxide, however, combines with -oxygen under suitable conditions at a temperature not exceeding 450 -deg. C. and forms a higher oxide, and it is probable that this unites -with the lime to form calcium plumbate. The reaction between sulphides -and lime when intimately mixed and heated may be put down as - - CaO + PbS = CaS + PbO. - -In contact with the air the calcium sulphide oxidizes to sulphite, then -to sulphate, then reacts with lead oxide, giving calcium plumbate and -sulphur dioxide, - - CaSO₄ + PbO = CaPbO₃ + SO₂. - -Further, calcium sulphate will also react with galena, giving calcium -sulphide and lead sulphate; the calcium sulphide is oxidized, by air -blown through, to calcium sulphate again, the ultimate reaction being - - CaSO₄ + PbS + O = CaPbO₃ + SO₂. - -In all cases the action is oxidizing and desulphurizing. It was found -that oxides of iron and manganese will, to a certain extent, serve the -same purpose as lime, and on application to complex ores, especially -those containing much blende, that these may be desulphurized as well -as galena. In the case of zinc sulphide the decomposition is probably -due to the interaction of sulphide and sulphate. - - ZnS + 3ZnSO₄ = 4ZnO + 4SO₂. - -The process has now been adopted by the Broken Hill Proprietary -Company at its works at Port Pirie, the Tasmanian Smelting Company, -Zeehan, the Fremantle Smelting Works, West Australia, and the Sulphide -Corporation’s works at Cockle Creek, New South Wales. - -The operations carried on at the Tasmania Smelting Works comprise -mixing pulverized limestone, galena and slag-making materials and -introducing the mixture either into hand-rabbled reverberatories or -mechanical furnaces with rotating hearths. After a roast, during -which the materials have become well mixed and most of the limestone -converted into sulphate and about half of the sulphur expelled, the -granular product is run while still hot into the Huntington-Heberlein -converters. These consist of inverted sheet-iron cones, hung on -trunnions, the diameter being 5 ft. 6 in. and the depth 5 ft. A -perforated plate or colander is placed as a diaphragm across the apex -of the cone, the small conical space below serving as a wind-box into -which compressed air is forced. A hood above the converter serves to -carry away waste gases. As soon as the vessel is filled, air under a -pressure of 17 oz. is forced through the mass, which rapidly warms up, -giving off sulphur dioxide abundantly. The temperature rises and the -mixture fuses, and in from two to four hours the action is complete. -The sulphur is reduced from 10 to 1 per cent., and the whole mass is -fritted and fused together. The converter is emptied by inverting it, -when the sintered mass falls out and is broken up and sent to the -smelters. There are 12 converters, of the size indicated, for the two -mechanical furnaces, of 15 ft. diameter. Larger converters of the -same type were erected to deal with the product from the hand-rabbled -roasters. - -At Cockle Creek, New South Wales, the galena concentrate is reduced -to 1.5 mm., more than 60 per cent. of the material being finer; the -limestone is crushed down to from 10 to 16 mesh; silica is also added, -if it does not exist in the ore, so that, excluding the lead, the rest -of the bases will be in such proportion as to form a slag running about -20 per cent. silica. The mixture may contain from 25 to 50 per cent. -lead, and from 6 to 9 per cent. lime; if too much lime is added the -final product is powdery, instead of being in a fused condition. This -is given a preliminary roast in a Godfrey furnace. - -The Godfrey furnace is characterized by a rotating, circular hearth -and a low dome-shaped roof. Ore is fed through a hopper at the center -and deflected outward by blades attached to a fixed radial arm. At -each revolution the ore is turned over and moved outward, the mount of -deflection of the blades, which are adjustable, and rate of rotation of -the hearth, determining the output. - -The hot semi-roasted ore is discharged through a slot at the -circumference of the roaster. This may contain from 12 to 6.5 per -cent. of sulphur, but from 6.5 to 8 per cent. is held to be the most -suitable quantity for the subsequent operations. Thorough mixing is of -the utmost importance, for if this is not done the mass will “volcano” -in the converter; that is, channels will form in the mass through which -the gases will escape, leaving lumps of untouched material alongside. -The action can be started if a little red-hot ore is run into the -converter and cold ore placed above it; the whole mass will become -heated up, and the products will fuse, and sinter into a homogeneous -mass showing none of the original ingredients. At Cockle Creek the time -taken is stated to be five hours; a small air-pressure is turned on at -first, and ultimately it is increased to 20 oz. - -Operations at Port Pirie are conducted on a much larger scale. A -mixture of pulverized galena, powdery limestone, ironstone and sand -is fed into Ropp furnaces, of which there are five, by means of a -fluted roll placed at the base of a hopper. Each roaster deals with -100 tons of the mixture in 24 hours. About 50 per cent. of the sulphur -is eliminated from the ore by the Ropps (the galena in this case being -admixed with a large amount of blende, there being only 55 per cent. -of lead and 10 per cent. of zinc in the concentrate produced at the -Proprietary mine). The hot ore from the roasters is trucked to the -converters, there being 17 of these ranged in line. The converters here -are large segmental cast-iron pots hung on trunnions; each is about 8 -ft. diameter and 6 ft. deep, and holds an 8-ton charge. At about two -feet from the bottom an annular perforated plate fits horizontally; -a shallow frustrum of a cone, also perforated, rests on this; while -a plate with a few perforations closes the top of the frustrum. The -whole serves as a wind-box. A conical hood with flanged edges rests -on the flanged edges of the converter, giving a close joint. This -hood is provided with doors which allow the charge to be barred if -necessary. A pipe about 1 ft. 9 in. diameter, fitted with a telescopic -sliding arrangement, allows for the raising or lowering of the hood by -block and tackle, and thus enables the converter to be tilted up and -its products emptied. The cast-iron pots stand very well; they crack -sometimes, but they can be patched up with an iron strap and rivets. -Only two pots have been lost in 18 months. - -Air enters at a pressure of about 24 oz. and the time taken for -conversion is about four hours. The sulphur contents are reduced to -about three per cent. It is found that the top of the charge is not so -well converted as the interior. There is practically no loss of lead -or silver due to volatilization and very little due to escape of zinc. -It has also been found that practically all the limestone fed into -the Ropp is converted into calcium sulphate; also that a considerable -portion of lead becomes sulphate, and it is considered that lead -sulphate is as necessary for the process as galena. - -The value of the process may be judged from the fact that better work -is now done with 8 blast furnaces than was done with 13 before the -process was adopted. In addition to the sintered product from the -Huntington-Heberlein pots, sintered slime, obtained by heap roasting, -and flux consisting of limestone and ironstone, are fed into the -furnaces, which take 2000 long tons per day of ore, fluxes and fuel. -The slags now being produced average: SiO₂, 25 to 26 per cent.; FeO, -1 to 3 per cent.; MnO, 5 to 5.5; CaO, 15.5 to 17; ZnO, 13; Al₂O₃, -6.5; S, 3 to 4; Pb, by wet assay, 1.2 to 1.5 per cent.; and Ag, 0.7 oz. -per ton. Although this comparatively large quantity of sulphur remains, -yet no matte is formed. - - - - - THE HUNTINGTON-HEBERLEIN PROCESS AT FRIEDRICHSHÜTTE[29] - - BY A. BIERNBAUM - - (September 2, 1905) - - -Nothing, for some time past, has caused such a stir in the -metallurgical treatment of lead ores, and produced such radical -changes at many lead smelting works, as the introduction of the -Huntington-Heberlein process. This process (which it may be remarked, -incidentally, has given rise to the invention of several similar -processes) represents an important advance in lead smelting, and, -now that it has been in use for some time at the Friedrichshütte, -near Tarnowitz, in Upper Silesia, and has there undergone further -improvement in several respects, a comparison of this process with the -earlier roasting process is of interest. - -At the above-mentioned works, up to 1900 the lead ore was -treated exclusively (1) by smelting in reverberatory furnaces -(Tarnowitzeröfen), and (2) by roasting in reverberatory-sintering -furnaces roasted material in the shaft furnace. The factor which -determined whether the treatment was to be effected in the -reverberatory-smelting or in the roasting-sintering furnace was the -percentage of lead and zinc in the ores; those comparatively rich in -lead and poor in zinc being worked up in the former, with partial -production of pig-lead; while those poorer in lead and richer in zinc -were treated in the latter. About two-fifths of the lead ores annually -worked up were charged into the reverberatory-smelting furnaces, and -three-fifths into the sintering furnaces. - -In 1900 there were available 10 reverberatory-smelting and nine -sintering furnaces. These were worked exclusively by hand. - -The sintered product of the roasting furnaces, and the gray slag from -the reverberatory-smelting furnaces, were transferred to the shaft -furnaces for further treatment, and were therein smelted together with -the requisite fluxes. Eight such furnaces (8 m. high, and 1.4 m., 1.6 -m., and 1.8 m. respectively in diameter at the tuyeres), partly with -three and partly with five or eight tuyeres, were at that time in use. - -Now that the Huntington-Heberlein process has been completely -installed, the reverberatory-smelting furnaces have been shut down -entirely, and the sintering furnaces also for the most part; all -kinds of lead ore, with a single exception, are worked up by the -Huntington-Heberlein process, irrespective of the contents of lead and -zinc. An exceedingly small proportion of the ore treated, viz., the -low-grade concentrate (Herdschlieche) containing 25 to 35 per cent. Pb, -is still roasted in the old sintering furnace, together with various -between-products (such as dust, fume, scaffoldings, and matte); these -are scorified by the aid of the high percentage of silica in the -material. - -For roasting lead ores at the present time there are six round -mechanical roasters of 6 m. diameter, one of 8 m. diameter, and two -ordinary, stationary Huntington-Heberlein furnaces. The latter (which -represent the primitive Huntington-Heberlein furnaces, requiring manual -labor) have recently been shut down, and will probably never be used -again. In the mechanical Huntington-Heberlein furnace, roasting of lead -ore is carried only to such a point that a small portion of the lead -sulphide is converted into sulphate. The desulphurization of the ore -is completed in the so-called converter (made of iron, pear-shaped or -hemispherical in form) in which the charge, up to this stage loosely -mixed, is blown to a solid mass. - -Owing to the ready fusibility of this product (which still contains, -as a rule, up to 1.5 per cent. sulphur as sulphide), it is possible to -use shaft furnaces of rather large dimensions; therefore a round shaft -furnace (2.4 m. diameter at the tuyeres, 7 m. high, and furnished with -15 tuyeres) was built. In this furnace nearly the whole of the roasted -ore from the Huntington-Heberlein converters is now smelted, some of -the smaller shaft furnaces being used occasionally. The introduction -of the new process has caused no noteworthy change in the subsequent -treatment of the work-lead. - -In the following study I shall discuss the treatment of a given annual -quantity of ore (50,000 tons), which is the actual figure at the -Friedrichshütte at the present time. - -1. _Roasting Furnaces._—A reverberatory-smelting furnace used to treat -5 tons of ore in 24 hours; a roasting-sintering furnace, 8 tons. -Assuming the ratios previously stated, the annual treatment by the -former process would be 20,000 tons, and by the latter 30,000 tons. -On the basis of 300 working days per year, and no prolonged stoppages -for furnace repairs (though considering the high temperatures of these -furnaces this record would hardly be expected), there would be required: - - 20,000 ÷ (5 × 300) = 13.3 (or 13 to 14 reverberatory furnaces). - 30,000 ÷ (8 × 300) = 12.5 (or 12 to 13 sintering furnaces). - -The capacity of a stationary Huntington-Heberlein furnace is 18 tons; -hence in order to treat the same quantity of ores there would be -required: - - 50,000 ÷ (18 × 300) = 9.3 (or 9 to 10 Huntington-Heberlein furnaces). - -With the revolving-hearth roasters (of 6 m. diameter) working a total -charge of at least 27 tons of ore, there would be required: - - 50,000 ÷ (27 × 300) = 6.1 (or 6 to 7 roasters). - -Still better results are obtained with the 8 m. round roaster, which -has been in operation for some time; in this, 55 tons of ore can be -roasted daily. Three such furnaces would therefore suffice for working -up the whole of the ore charged per annum. - -Now, making due provision for reserve furnaces, to work up 50,000 tons -of ore would require: - - Reverberatory (15) and sintering furnaces (15) 30 - Stationary Huntington-Heberlein furnaces 12 - 6 m. revolving-hearth furnaces 8 - 8 m. revolving-hearth furnaces 4 - -Similar relations hold good regarding the number of workmen -attending the furnaces, there being required, daily, six men for the -reverberatory furnace; eight men for the sintering furnace; ten men for -the stationary; and six men for the mechanical Huntington-Heberlein -furnace; or, for 14 reverberatory furnaces, daily, 84 men; for -sintering furnaces, daily, 104 men; total, 188 men. While for 10 -stationary Huntington-Heberlein furnaces, 100 men are required; and -for 7 mechanical Huntington-Heberlein furnaces, daily, 42 men. It is -expected that only 14 men (working in two shifts) will be required to -run the new installation with 8 m. round roasters. - -It is true that the exclusion of human labor here has been carried to -an extreme. The roasters and converters will be charged exclusively -by mechanical means; thus every contact of the workmen with the -lead-containing material is avoided until the treatment of the roasted -material in the converters is completed. - -From the data given above, the capacity of each individual workman -is readily determined, as follows: With the reverberatory-smelting -furnace, each man daily works up 0.83 tons; with the sintering furnace, -1 ton; with the stationary Huntington-Heberlein furnace, 1.8 tons; -with the 6 m. revolving-hearth furnace, 4.5 tons; and with the 8 m. -revolving-hearth furnace, 11.8 tons. - -A significant change has also taken place in coal consumption. Thus, -when working with the reverberatory and sintering furnaces in order to -attain the requisite temperature of 1000 deg. C., there was required -not only a comparatively high-grade coal, but also a large quantity of -it. A reverberatory furnace consumed about 503 kg., a sintering furnace -about 287 kg., of coal per ton of ore. For roasting the ore in the -stationary and also in the mechanical Huntington-Heberlein furnaces, a -lower temperature (at most 700 deg. C.) is sufficient, as the roasting -proper of the ore is effected in the converters, and the sulphur -furnishes the actual fuel. For this reason, the consumption of coal is -much lower. The comparative figures per ton of ore are as follows: In -the reverberatory furnace, 50.3 per cent.; in the sintering furnace, -28.7 per cent.; in the stationary Huntington-Heberlein furnace, 10.3 -per cent.; and in the Huntington-Heberlein revolving-hearth furnace, -7.3 per cent. - -But there is another technical advantage of the Huntington-Heberlein -process which should be mentioned. It is well known that the -volatilization of lead at high temperatures is an exceedingly -troublesome factor in the running of a lead-smelting plant; the -recovery of the valuable fume is difficult, and requires condensing -apparatus, to say nothing of the unhealthful character of the volatile -lead compounds. This volatilization is of course particularly marked at -the high temperatures employed when working with reverberatory-smelting -furnaces; the same is true, in a somewhat less degree, of the sintering -furnaces. In consequence of the markedly lower temperature to which -the charge is heated in the Huntington-Heberlein furnace, and also of -the peculiar mode of completing the roast in blast-converters, the -production of fume is so reduced that the difference between the values -recovered in the old and the new processes is very striking. Whereas, -in 1900, in working up 12,922 tons of ore in the reverberatory-smelting -furnace, and 14,497 tons in the sintering furnace (27,419 tons in -all), there was recovered 2470 tons (or 9 per cent.) as fume from -the condensers and smoke flues, the quantity of fume recovered, in -1903, fell to 879 tons (or 1.8 per cent.), out of the 48,208 tons of -ore roasted, and this notwithstanding the fact that in the meantime -fume-condensing appliances had been considerably expanded and improved, -whereby the collection was much more efficient. - -Lastly, the zinc content of the ores no longer exerts the same -unfavorable influence as in the old process (wherein it was advisable -to subject ore containing much blende to a final washing before -proceeding to the actual metallurgical treatment). In the new process, -the ores are simply roasted without regard to their zinc content. In -this connection it has been found that a considerable proportion of the -zinc passes off with the fume, and that the roasted material usually -contains a quantity of zinc so small that it no longer causes any -trouble in the shaft furnace. It may also be mentioned here that the -ore-dressing plants recently installed in the mines of Upper Silesia -have resulted in a more perfect separation of the blende. - -_Shaft Furnaces._—The finished product from the Huntington-Heberlein -blast-converters is of a porous character, and already contains a -part of the flux materials (such as limestone, silica and iron) which -are required for the shaft-furnace charge. It is just these two -characteristics of the roasted product (its porous nature, on the one -hand, leading to its more perfect reduction by the furnace gases; and, -on the other hand, the admixture of fluxes in the molten condition, -resulting in a more complete utilization of the temperature), which, -together with its higher lead and lower zinc content, determine its -ready fusibility. If we further consider that it is possible in the new -process to make the total charge of the shaft furnace richer in lead -than formerly (two-thirds of the total charge as against one-third), -and that a higher blast pressure can be used without danger, it follows -immediately that the capacity of a shaft furnace is much greater by -the new process than by the old method of working. The daily production -of the shaft furnaces on the old and the new process is as shown in the -table given herewith: - - ─────────────┬─────────────────────────┬─────────┬──────────────────── - │ │ CHARGE │ WORK-LEAD - TYPE OF SHAFT│ CHARACTER OF CHARGE │ PER DAY,│ PRODUCED - FURNACE │ │ TONS │ PER DAY, TONS - ─────────────┼─────────────────────────┼─────────┼──────────────────── - 3 tuyeres │{ Gray slag from } │ 36 │ 6 to 7 } - │{ reverberatory } │ │ } - │{ furnaces and } │ │ } Low- - │{ sintered concentrate } │ │ }pressure - │ │ │ } Blast - 8 tuyeres │ ” ” │ 36 to 38│ 6 to 8 } - │ │ │ } - 3 tuyeres │{ Roasted product of } │ 36 │ 11 to 12 } - │{ Huntington-Heberlein } │ │ - │{ process } │ │ - │ │ │ - 8 tuyeres │ ” ” │ 65 to 72│ 24 to 26 } High- - │ │ │ }pressure - 15 tuyeres │ ” ” │ 270 │ 90 to 100 } Blast - ─────────────┴─────────────────────────┴─────────┴──────────────────── - -It should be noted that the figure given for the furnace with 15 -tuyeres represents the average for 1904; this average is lowered by the -circumstance that during this period there was frequently a deficiency -of roasted material, and the furnace had to work with low-pressure -blast. A truer impression can be gained from the month of March, 1905, -for instance, during which time this furnace worked under normal -conditions; the results are as follows: - -The average for March, 1905, was: Ore charged, 8,269.715 tons; coke, -652.441 tons; total, 8,922.156 tons. Or, in 24 hours: Ore charged, -266.765 tons; coke, 21.046 tons; total, 287.811 tons. The production of -work-lead was 3,133.245 tons, or 101.069 tons per day. - -The maximum production of roasted ore was 210 tons, on June 30, 1905, -when the total charge was: Ore, 327.38 tons; coke, 25.2 tons; total, -352.58 tons. The quantity of work-lead produced on that day was 120.695 -tons, while the largest quantity previously produced in one day was -124.86 tons. It should also be mentioned that the lead tenor of the -slag is almost invariably below 1 per cent.; it usually lies between -0.3 and 0.5 per cent. - -As in the case of the roasting furnaces, the productive capacity of -the shaft furnace also comes out clearly if we figure the number -of furnaces required, on the basis of an annual consumption of -50,000 tons of ore. If we consider 1 ton of the roasted material as -equivalent to 1 ton of ore (which is about right in the case of the -Huntington-Heberlein material, but is rather a high estimate in the -case of the product of the sintering furnace), then, in the old process -(where one-third of the charge was lead-bearing material), 12 tons -could be smelted daily. There would therefore be needed at least: - - 50,000 ÷ (12 × 300) = 14 three-tuyere shaft furnaces. - -Since, as already mentioned, the lead-bearing part of the charge -constitutes two-thirds of the whole in the Huntington-Heberlein -process, the number of shaft furnaces of different types, as compared -with the foregoing, would figure out: - - 3-tuyere shaft furnace, with product of sintering furnace, - 50,000 ÷ (12 × 300) = 14 furnaces; - - 3-tuyere shaft furnace, with product of Huntington-Heberlein furnace, - 50,000 ÷ (24 × 300) = 7 furnaces; - - 8-tuyere shaft furnace, with product of Huntington-Heberlein furnace, - 50,000 ÷ (48 × 300) = 3.4 (say 4) furnaces; - - 15-tuyere shaft furnace, with product of Huntington-Heberlein furnace, - 50,000 ÷ (180 × 300) = 1 furnace. - -Running regularly and without interruption, the large shaft furnace is -therefore fully capable of coping with the Huntington-Heberlein roasted -material at the present rate of production. - -As regards the number of workmen and the product turned out per man, -no such marked difference is produced by the introduction of the -Huntington-Heberlein process in the case of the shaft furnace as there -was noted for the roasting operation. This is chiefly due to the fact -that the work which requires the more power (such as charging of the -furnaces, conveying away the slag and pouring the lead) can be executed -only in part by mechanical means. Nevertheless, it will be seen from -the table given herewith that, on the one hand, the number of men -required for the charge worked up is smaller; and, on the other, the -product turned out per man has risen somewhat. - - ─────────┬─────────┬────────┬──────────┬────────┬─────────────┬─────── - TYPE OF │CHARACTER│ CHARGE │NUMBER OF │ CHARGE │DAILY OUTPUT │OUTPUT - SHAFT │OF CHARGE│PER DAY,│FURNACEMEN│PER MAN,│OF WORK-LEAD,│PER MAN, - FURNACE │ │ TONS │ │ TONS │ TONS │ TONS - ─────────┼─────────┼────────┼──────────┼────────┼─────────────┼─────── - 3 tuyere│ A │ 36 │ 6 │ 6.0 │ 6 │ 1.0 - 8 tuyere│ B │ 38 │ 6 │ 6.3 │ 8 │ 1.3 - 3 tuyere│ C │ 36 │ 6 │ 6.0 │ 12 │ 2.0 - 8 tuyere│ D │ 72 │ 12 │ 6.0 │ 26 │ 2.1 - 15 tuyere│ E │ 270 │ 34 │ 7.9 │ 90 │ 2.6 - ─────────┴─────────┴────────┴──────────┴────────┴─────────────┴─────── - - ┌──────────┬──────────────────────────────────────┐ - │ CHARACTER│ CHARACTER OF CHARGE │ - │ OF CHARGE│ │ - │ CODE │ │ - ├──────────┼──────────────────────────────────────┤ - │ A │ Sintered concentrate and gray slag │ - │ B │ from reverberatory furnace. │ - │ B │ Gray slag from reverberatory furnace.│ - │ C │ Huntington-Heberlein product. │ - │ D │ Huntington-Heberlein product. │ - │ E │ Huntington-Heberlein product. │ - └──────────┴──────────────────────────────────────┘ - -A slight difference only is produced by the new process in the -consumption of coke; the economy is a little over 1 per cent., the -coke consumed being reduced from 9.39 per cent. to 8.17 per cent. of -the total charge. But with the high price of coke, even this small -difference represents a considerable lowering of the cost of production. - -With the great increase in the blast pressure, it would be supposed -that the losses in fume would be much greater than with the former -method of working. But this is not the case; on the contrary, all -experience so far shows that there is much less fume developed. In -1904, for instance, the shaft-furnace fume recovered in the condensing -system amounted to only 1.06 per cent. of the roasted material, or -0.64 per cent. of the total charge, as against 2.03 and 1.0 per cent., -respectively, in former years. The observations made on the quantity of -flue dust carried away with the gases escaping into the air through the -stack showed that it is almost nil. - -Now, from the loss in fume being slight, from the tenor of lead in the -slag being low, and, on the one hand, from the quantity of lead-matte -produced being much less than before, while on the other the losses in -roasting the ore are greatly reduced—from all these considerations, it -is clear that the total yield must have been much improved. As a matter -of fact, the yield of lead and silver has been increased by at least 6 -to 8 per cent. - -_Economic Results._—As regards the economical value of the new process, -for obvious reasons no data can be furnished of the exact expenditure, -i.e., the actual total cost of roasting and smelting the ore. But -this at least is placed beyond doubt by what has been developed above, -namely, that considerable saving must be effected in the roasting, -and especially in the smelting, as compared with the former mode of -working. If we take into account only the economy which is gained -in wages through the increase in the material which one workman can -handle, and that resulting from the reduced consumption of coal and -coke, these alone will show sufficiently that an important diminution -of working cost has taken place. The objection which might be raised, -that the saving effected by reducing manual labor may be neutralized -by the expense of mechanical power (actuating the roasters, furnishing -the compressed blast, etc.), cannot be regarded as justified, as the -cost of mechanical work is comparatively low. Thus, for instance, the -large 8 m. furnace and the small, round furnaces require 15 h.p. if -worked by electricity. According to an exact calculation, the cost -(to the producer) of the h.p. hour, inclusive of machinery, figures -out to 3.6 pfennigs (0.9c.); hence the daily expense for running the -revolving-hearth furnaces amounts to: 15 × 3.6 pfg. × 24 = 12.96 marks -($3.42). As the seven furnaces together work up: (6 × 27) + 55 = 217 -tons of ore, the cost per ton of ore is about 0.06 mark (1.5c.). - -The requisite blast is produced by means of single-compression Encke -blowers, of which one is quite sufficient when running at full load, -and then consumes 34 h.p. The daily expenses are accordingly: 34 × 3.6 -pfg. × 24 = 29.28 marks ($7.32); or per ton of ore, 29.28 ÷ 217 = 0.14 -mark (3.5c.). Therefore the total expense for the mechanical work in -roasting the ore amounts to 0.06 + 0.14 = 0.20 mark (5c.). - -However, the cost of roasting is much more affected by the expense -for keeping the furnaces in repair; another important factor is the -acquisition and maintenance of the tools. Both in the case of the -sintering and also the reverberatory-smelting furnace, the cost of -keeping in repair was high; the consumption of iron was especially -large, owing to the rapid wear of the tools. This was not surprising, -considering that a notably higher temperature prevailed in the -reverberatory and sintering furnaces than in the new roasters, in which -the temperature strictly ought not to rise above 700 deg. C. But in the -old type of furnace the high temperature and the constant working with -the iron tools caused their rapid wear, thus creating a large item for -iron and steel and smith work. In the new process (and more especially -in the revolving-hearth roasters) this disadvantage does not arise. In -this case there is practically no work on the furnace, and the wear -and tear of iron is small. Also, the cost of keeping the furnaces -in repair when working regularly is small as compared with the old -process. In the year 1900, for instance, the cost of maintenance and -tools for the reverberatory and sintering furnaces came to 20,701.93 -marks ($5,175.48) for treating 27,419.75 tons of ore. Per ton of ore, -this represents 0.75 mark (19c.). In the year 1903, on the other -hand, only 9,074.17 marks ($2,268.54) were expended, although 48,208 -tons of ore were worked up in the three stationary and six mechanical -Huntington-Heberlein furnaces. The cost of maintenance was, therefore, -in this case 0.18 mark (4.5c.) per ton of ore. - -In the cost of smelting in the shaft furnace, only a slight difference -in favor of the Huntington-Heberlein process is found if the estimate -is based on the total charge; but a marked difference is shown if it is -referred to the lead-bearing portion of the charge, or to the work-lead -produced. Thus the cost of maintenance and total cost of smelting, -figured for one ton of ore, without taking into account general -expenses, have been tabulated as follows: - - ────────────────────────────┬──────────────────────────────── - │REDUCTION IN EXPENSES PER TON OF - ├────────┬──────────┬──────────── - │ TOTAL │ LEAD ORE │ WORK-LEAD - │ CHARGE │ │ - ────────────────────────────┼────────┼──────────┼──────────── - (_a_) Cost of maintenance │ 0.01M │ 0.38M │ 0.67M - │(0.25c) │ (9.5c) │ (16.75c) - │ │ │ - (_b_) Total cost of smelting│ 0.20M │ 6.46M │ 11.48M - │ (5c) │ ($1.615) │ ($2.87) - ────────────────────────────┴────────┴──────────┴──────────── - -The marked reduction in the expenses, as referred to the lead-ore and -the work-lead produced, is determined (as was pointed out above) by the -greater lead content of the charge, and by the larger yield of lead -consequent thereon. The advantage of longer smelting campaigns (which -ultimately were mostly prolonged to one year) also makes itself felt; -it would be still more marked, if the shaft furnace (which was still in -working condition after it was blown out) had been run on for some time -longer. - -Finally, if we examine the question of the space taken up by the plant -(which, owing to the scarcity of suitably located building sites, -would have been important at the Friedrichshütte at the time when the -quantity of ore treated was suddenly doubled), here again we shall -recognize the great advantage which this establishment has gained from -the Huntington-Heberlein process. - -As was calculated above, there would have been required 15 -reverberatory and 15 sintering furnaces to cope with the quantity of -ore treated. As a reverberatory requires, in round numbers, 120 sq. m. -(1290 sq. ft.), and a sintering furnace 200 sq. m. (2153 sq. ft.); and -as fully 100 sq. m. (1080 sq. ft.) must be allowed for each furnace for -a dumping ground, therefore the 15 reverberatory furnaces would have -required an area of 15 × 120 + 15 × 100 = 3300 sq. m.; the 15 sintering -furnaces would have required 15 × 200 + 15 × 100 = 4500 sq. m.; in -all 3300 + 4500 = 7800 sq. m. (83,960 sq. ft.). The 12 stationary -Huntington-Heberlein furnaces (built together two and two) would take -up a space of 6 × 200 + 12 × 100 = 2400 sq. m. (25,830 sq. ft.). -Similarly, 8 small furnaces would require 8 × 100 + 8 × 100 = 1600 sq. -m. (17,222 sq. ft.); while for the new installation of four 8-meter -revolving-hearth furnaces and 10 large converters, only 1320 sq. m. -(14,120 sq. ft.) have been allowed. - -For shaft furnaces with three or eight tuyeres, which were run with -low-pressure blast for the material roasted on the old plan, the total -area built upon was 18 × 16.5 = 297 sq. m.; while a further area of 18 -× 14 = 250 sq. m. was hitherto provided, and was found sufficient for -dumping slag when working regularly. Therefore, the installation of -shaft furnaces formerly in existence, after requisite enlargement to -14 furnaces, would have demanded a space of 7 × 297 + 7 × 250 = 3829 -sq. m. (42,215 sq. ft.). If four of the small shaft furnaces had been -reconstructed for eight tuyeres, and run with Huntington-Heberlein -roasted material, using high-pressure blast, the area occupied would -have been reduced to 2 × 297 + 2 × 250 sq. m. = 1094 sq. m. (11,776 sq. -ft.). - -Still more favorable are the conditions of area required in the case of -the large shaft furnace. This furnace stands in a building covering an -area of 350 sq. m. (3767 sq. ft.), which is more than sufficient room. -The slag-yard (situated in front of this building, and amply large -enough for 36 hours’ run) has an area of 250 sq. m. (2691 sq. ft.); -thus the space occupied by the large shaft furnace, including a yard of -170 sq. m. (1830 sq. ft.), is in all 780 sq. m. (8396 sq. ft.). - -After completion of the new roasting plant and the large shaft furnace -in connection with it, there would be occupied 1320 + 780 = 2100 sq. -m. (2260 sq. ft.); and if the system of reverberatory and sintering -furnaces had been continued (with the requisite additions thereto and -to the old shaft-furnace system), there would have been required 11,629 -sq. m. (125,214 sq. ft.). In the estimate above given no regard has -been paid to any of the auxiliary installations (dust chambers, etc.), -which, just as in the case of the old process, would have had to be -provided on a large scale. - -It is of course self-evident that both the principal and the auxiliary -installations in the old process would not only have involved a high -first cost, but would also, on account of their extensive dimensions, -have caused considerably greater annual expense for maintenance. - - - - - THE HUNTINGTON-HEBERLEIN PROCESS FROM THE HYGIENIC STANDPOINT[30] - - BY A. BIERNBAUM - - (October 14, 1905) - - -With regard to the hygienic improvements which the Huntington-Heberlein -process offers, we must first deal with the questions: What were -the sources of danger in the old process, and in what way are these -now diminished or eliminated? The only danger which enters into -consideration is lead-poisoning, other influences detrimental to health -being the same in one process as the other. - -With the reverberatory-smelting and roasting-sintering furnaces, the -chief danger of lead-poisoning lies in the metallic vapor evolved -during the withdrawal of the roasted charge from the furnace. It is -true that appliances may be provided, by which these vapors are drawn -off or led back into the furnace during this operation; but, even -working with utmost care, it is impossible to insure the complete -elimination of lead fumes, especially in wheeling away the pots -filled with the red-hot sintered product. Moreover, the work at the -reverberatory-smelting and roasting-sintering furnaces involves great -physical exertion, wherefore the respiratory organs of the workmen -are stimulated to full activity, while the exposure to the intense -heat causes the men to perspire freely. Hence, as has been established -medically, the absorption of the poisonous metallic compounds (which -are partially soluble in the perspiration) into the system is favored -both by inhalation of the lead vapor and by its penetration into the -pores of the skin, opened by the perspiration. - -A further danger of lead-poisoning was occasioned by the frequently -recurring work of clearing out the dust flues. The smoke from the -reverberatory-smelting furnace especially contained oxidized lead -compounds, which on absorption into the human body might readily be -dissolved by the acids of the stomach, and thus endanger the health of -the workmen. - -In the Huntington-Heberlein furnaces, on the other hand, although the -charge is raked forward and turned over by hand, it is not withdrawn, -as in the old furnaces, by an opening situated next to the fire, but -is emptied at a point opposite into the converters which are placed -in front of the furnace. Moreover, the converters are filled with the -charge at a much lower temperature. Inasmuch as this charge has already -cooled down considerably, there can be practically no volatilization of -lead. The small quantity of gas which may nevertheless be evolved is -drawn off by fans through hoods placed above the converters. - -A further improvement, from the hygienic point of view, is in the use -of the mechanical furnaces, from which the converters can be filled -automatically (almost without manual labor, and with absolute exclusion -of smoke). The converters are then placed on their stands and blown. -This work also is carried out under hoods, as gas-tight as possible, -furnished with a few closable working apertures. During the blowing -of the material, the work of the attendant consists solely in keeping -up the charge by adding more cold material and filling any holes that -may be formed. It does not entail nearly as much physical strain as -the handling of the heavy iron tools and the continued exposure of the -workmen to the hottest part of the furnace, which the former roasting -process involved. - -Some experiments carried out with larger converters (of 4 and 10 -ton capacity) have indicated the direction in which the advantages -mentioned above may probably be developed to such a point that the -danger of lead-poisoning need hardly enter into consideration. Both -the charging of the revolving-hearth furnaces and the filling of the -converters are to be effected mechanically. Furthermore, in the case -of the large converters the filling up of holes becomes unnecessary, -and no manual work of any kind is required during the whole time -of blowing. The converters can be so perfectly enclosed in hoods -that the escape of gases into the working-rooms becomes impossible, -and lead-poisoning of the men can occur only under quite unusual -circumstances. - -The beneficial influence on the health of the workmen attending -on the roasting furnaces, occasioned by the introduction of the -Huntington-Heberlein process, can be seen from the statistics of -sickness from lead-poisoning for the years 1902 to 1904, as given -herewith: - - ─────────┬──────┬──────┬──────────────────────────────┬─────────────── - │ │ │ LEAD-POISONING │ CASES - │ │ ├─────────────┬────────────────┤ CONTRACTED - │ │ │NO. OF CASES │DAYS OF SICKNESS│AT REVER.│ AT - ─────────┼──────┼──────┼─────┬───────┼───────┬────────┤ AND │H. H. SICKNESS - METHOD OF│ YEAR │NO. OF│TOTAL│PER 100│ TOTAL │PER 100 │ SINT. │ FUR. - WORKING │ │ MEN │ │PERSONS│ │PERSONS │ FUR. │ - ─────────┼──────┼──────┼─────┼───────┼───────┼────────┼─────────┼───── - │ │ │ │ │ │ │ │ - Old │{ 1902│ 93 │ 15 │ 16.1 │ 246 │ 264.5 │ 11 │ 4 - │{ 1903│ 86 │ 12 │ 13.9 │ 222 │ 258.1 │ 7 │ 5 - │ │ │ │ │ │ │ │ - H.-H. │ 1904│ 87 │ 8 │ 9.2 │ 242 │ 278.2 │ 6 │ 2 - ─────────┴──────┴──────┴─────┴───────┴───────┴────────┴─────────┴───── - -This shows a gratifying decrease in the number of cases, namely, from -16.1 to 9.2 per cent.; this decrease would have been still greater if -Huntington-Heberlein furnaces had been in use exclusively. However, -most of the time two or three sintering furnaces were fired for -working up by-products, 16 to 18 men being engaged on that work. The -Huntington-Heberlein furnaces alone (at which, in the year 1904, 69 men -in all were occupied) show only 2.9 per cent. of cases. That the number -of days of illness was not reduced is due to the fact that the cases -among the gang of men working at the sintering furnaces were mostly of -long standing and took some time to cure. - -The noxious effects upon the health of the workmen in running the -shaft furnaces are due to the fumes from the products made in this -operation, such as work-lead, matte and slag, which flow out of the -furnace at a temperature far above their melting points. Even with -the old method of running the shaft furnaces the endeavor has always -been to provide as efficiently as possible against the danger caused -by this volatilization, and, wherever feasible, to install safety -appliances to prevent the escape of lead vapors into the work-rooms; -but these measures could not be made as thorough as in the case of the -Huntington-Heberlein process. - -The principal work in running the shaft furnaces, aside from the -charging, consists in tapping the slag and pouring out the work-lead. -Other unpleasant jobs are the barring down (which in the old process -had to be done frequently) and the cleaning out of the furnace after -blowing out. - -In the old process the slag formed in the furnace flows out -continuously through the tap-hole into iron pots placed in front of -the spout. A number of such pots are so arranged on a revolving table -that as soon as one is filled the next empty can be brought up to the -duct; thus the slag first poured in has time to cease fuming and to -solidify before it is removed. The vapors arising from the slag as it -flows out are conveyed away through hoods. At the same time with the -slag, lead matte also issues from the furnace. Now the greater the -quantity of lead matte, the more smoke is also produced; and, with -the comparatively high proportion of lead matte resulting from the -old process, the quantity of smoke was so great that the ventilation -appliances were no longer sufficient to cope with it, thus allowing -vapors to escape into the work-room. - -The work-lead collects at the back of the furnace in a well, from which -it is from time to time ladled into molds placed near by. If the lead -is allowed to cool sufficiently in the well, it does not fume much in -the ladling out. But when the furnace runs very hot (which sometimes -happens), the lead also is hotter and is more inclined to volatilize. -In this event the danger of lead-poisoning is very great, for the -workman has to stand near the lead sump. - -A still greater danger attends the work of barring down and cleaning -out the furnace. The barring down serves the purpose of loosening -the charge in the zone of fusion; at the same time it removes any -crusts formed on the sides of the furnace, or obstructions stopping -up the tuyeres. With the old furnaces, and their strong tendency to -crust, this work had to be undertaken almost every day, the men being -compelled to work for rather a long time and often very laboriously -with the heavy iron tools in the immediate neighborhood of the glowing -charge, the front of the furnace being torn open for this purpose. In -this operation they were exposed without protection to the metallic -vapors issuing from the furnace, inasmuch as the ventilating appliances -had to be partially removed during this time, in order to render it at -all possible to do the work. - -In a similar manner, but only at the time of shutting down a shaft -furnace, the cleaning out (that is to say, the withdrawing of no -longer fused but still red-hot portions of the charge left in the -furnace) is carried out. In this process, however, the glowing material -brought out could be quenched with cold water to such a point that the -evolution of metallic vapors could be largely avoided. - -Lastly, the mode of charging of the shaft furnace is also to be -regarded as a cause of poisoning, inasmuch as it is impossible to -avoid entirely the raising of dust in the repeated act of dumping and -turning over the materials for smelting, in preparing the mix, and in -subsequently charging the furnace. - -By the introduction of the Huntington-Heberlein process, all these -disadvantages, both in the roasting operation and in running the shaft -furnaces, are in part removed altogether, in part reduced to such a -degree that the danger of injury is brought to a minimum. - -In furnaces in which the product of the Huntington-Heberlein roast -is smelted, the slag is tapped only periodically at considerable -intervals; and, as there is less lead matte produced than formerly, the -quantity of smoke is never so great that the ventilating fan cannot -easily take care of it. There is therefore little chance of any smoke -escaping into the working-room. - -As the production of work-lead, especially in the case of the large -shaft furnace, is very considerable, so that the lead continually -flows out in a big stream into the well, the hand ladling has to -be abandoned. Therefore the lead is conducted to a large reservoir -standing near the sump, and is there allowed to cool below its -volatilizing temperature. As soon as this tank is full, the lead is -tapped off and (by the aid of a swinging gutter) is cast into molds -ready for this purpose. Both the sump and the reservoir-tank are placed -under a fume-hood. The swinging gutter is covered with sheet-iron lids -while tapping, so that any lead volatilized is conveyed by the gutter -itself to a hood attached to the reservoir; thus the escape of metallic -vapors into the working space is avoided, as far as possible. - -This method of pouring does not entail the same bodily exertion as the -ladling of the lead; moreover, as it requires but little time, it gives -the workmen frequent opportunity to rest. - -But one of the chief advantages of the Huntington-Heberlein process -lies in the entire omission of the barring down. If the running of the -shaft furnace is conducted with any degree of care, disorders in the -working of the furnace do not occur, and one can rely on a perfectly -regular course of the smelting process day after day. No formation -of any crusts interfering with the operation of the furnace has been -recorded during any of the campaigns, which have, in each case, lasted -nearly a year. - -As regards the cleaning out of the furnace, this cannot be avoided -on blowing out the Huntington-Heberlein shaft furnace; but at most -it occurs only once a year, and can be done with less danger to the -workmen, owing to the better equipment. - -Further, the charge is thrown straight into the furnace (in the case -of the large shaft furnace); thus the repeated turning over of the -smelting material, as formerly practised, becomes unnecessary, and the -deleterious influence of the unavoidable formation of dust is much -diminished. - -The accompanying statistics of sickness due to lead-poisoning in -connection with the operation of the shaft furnace (referring to the -same period of time as those given above for the roasting furnaces) -confirm the above statements. - - ────┬──────────┬──────────────────────────────────────────── - │ │ LEAD-POISONING—SHAFT FURNACES - │ ├─────────────────────┬────────────────────── - YEAR│NO. OF MEN│ CASES │ DAYS OF ILLNESS - │ ├─────┬───────────────┼─────┬──────────────── - │ │TOTAL│PER 100 PERSONS│TOTAL│PER 100 PERSONS - ────┼──────────┼─────┼───────────────┼─────┼──────────────── - 1902│ 250 │ 58 │ 23.2 │ 956 │ 382.4 - 1903│ 267 │ 59 │ 22.1 │1044 │ 391.0 - 1904│ 232 │ 24 │ 10.3 │ 530 │ 228.4 - ────┴──────────┴─────┴───────────────┴─────┴──────────────── - -If it were possible to make the necessary distinctions in the case of -the large shaft furnace, the diminution in sickness from lead-poisoning -would be still more apparent; for, among the furnace attendants proper, -there has been no illness; all cases of poisoning have occurred among -the men who prepare the charge, who break up the roasted material, and -others who are occupied with subsidiary work. Some of these are exposed -to illness through their own fault, owing to want of cleanliness, or to -neglect of every precautionary measure against lead-poisoning. - -Thus far we have dealt only with the advantages and improvements of the -Huntington-Heberlein process; we will now, in conclusion, consider also -its disadvantages. - -The chief drawback of the new process lies in the difficulty of -breaking up the blocks of the roasted product from the converters, a -labor which, apart from the great expense involved, is also unhealthy -for the workmen engaged thereon. Seemingly this evil is still further -increased by working with larger charges in the 10 ton converters, as -projected; but in this case it is proposed to place the converters in -an elevated position, and to cause the blocks to be shattered by their -fall from a certain hight, so that further breaking up will require -but little work. Trials made in this direction have already yielded -satisfactory results, and seem to promise that the disadvantage will in -time become less important. - -Another unpleasant feature is the presence (in the waste gases from the -converters) of a higher percentage of sulphur dioxide, the suppression -of which, if it is feasible at all, might be fraught with trouble and -expense. - -That the roaster gases from the reverberatory-smelting and sintering -furnaces did not show such a high percentage of sulphur dioxide must -be ascribed chiefly to the circumstance that the roasting was much -slower, and that the gases were largely diluted with air already at the -point where they are formed, as the work must always be done with the -working-doors open. In the Huntington-Heberlein process, on the other -hand, the aim is to prevent, as far as possible, the access of air from -outside while blowing the charge. The more perfectly this is effected, -and the greater the quantity of ore to be blown in the converters, the -higher will also be the percentage of sulphur dioxide in the waste -gases. This circumstance has not only furnished the inducement, but it -has rendered it possible to approach the plan of utilizing the sulphur -dioxide for the manufacture of sulphuric acid. If this should be done -successfully (which, according to the experiments carried out, there -is reasonable ground to expect), the present disadvantage might be -turned into an advantage. This has the more significance because an -essential constituent of the lead ore—the sulphur—will then no longer, -as hitherto, have to be regarded as wholly lost.[31] - - - - - THE HUNTINGTON-HEBERLEIN PROCESS - - BY THOMAS HUNTINGTON AND FERDINAND HEBERLEIN - - (May 26, 1906) - - -This process for roasting lead sulphide ores has now fairly -established itself in all parts of the world, and is recognized by -metallurgical engineers as a successful new departure in the method of -desulphurization. It offers the great advantage over previous methods -of being a more scientific application of the roasting reactions (of -the old well-used formulæ PbS + 3O = PbO + SO₂ and PbS + PbSO₄ -+ 2O = 2PbO + 2SO₂) and admits of larger quantities being handled -at a time, so that the use of fuel and labor are in proportion to the -results achieved, and also there is less waste all around in so far -as the factors necessary for the operation—fuel, labor and air—can -be more economically used. The workman’s time and strength are not -employed in laboriously shifting the ore from one part of the furnace -to another with a maximum amount of exertion and a minimum amount of -oxidation. The fuel consumed acts more directly upon the ore during the -first part of the process in the furnace and its place is taken by the -sulphur itself during the final and blowing stage, so that during the -whole series of operations more concentrated gases are produced and -consequently the large excess of heated air of the old processes is -avoided to such an extent that the gases can be used for the production -of sulphuric acid. - -With a modern well-constructed plant practically all the evils of -the old hand-roasting furnaces are avoided, and besides the notable -economy achieved by the H.-H. process itself, the health and well-being -of the workmen employed are greatly advanced, so that where hygienic -statistics are kept it is proved that lead-poisoning has greatly -diminished. It is only natural, therefore, that the H.-H. process -should have been a success from the start, popular alike with managers -and workmen once the difficulties inseparable from the introduction of -any new process were overcome. - -Simple as the process now appears, however, it is the result of many -years of study and experiment, not devoid of disappointments and at -times appearing to present a problem incapable of solution. The first -trials were made in the smelting works at Pertusola, Italy, as far -back as 1889, where considerable sums were devoted every year to this -experimental work and lead ore roasting was almost continuously on the -list of new work from 1875 on. - -It may be interesting to mention that at this time the Montevecchio -ores (containing about 70 per cent. lead and about 15 per cent. -sulphur, together with a certain amount of zinc and iron) were -considered highly refractory to roast, and the only ores approved of -by the management of the works at this date were the Monteponi and -San Giovanni first-class ores (containing about 80 per cent. lead), -and the second-class carbonates (with at least 60 per cent. lead and -5 per cent. sulphur). It must be noted that a modified Flintshire -reverberatory process was in use in the works, which could deal -satisfactorily only with this class of ore, so that, as these easy ores -diminished in quantity every year and their place was taken by the -“refractory” Montevecchio type, the roasting problem was always well to -the front at the Pertusola works. - -It may be asserted that almost every known method of desulphurization -was examined and experimented upon on a large scale. Gas firing was -exclusively used on certain classes of ores for several years with -considerable success, and revolving furnaces of the Brückner type—gas -fired—were also tried. Although varying degrees of success were -obtained, no really great progress was made in actual desulphurization; -methods were cheapened and larger quantities handled at a time, but -the final product—whether sintered or in a pulverulent state—seldom -averaged much under 5 per cent. sulphur, while the days of the -old “gray slags” (1 per cent. to 2 per cent. sulphur) from the -reverberatories totally disappeared, together with the class of ores -which produced them. - -During the long period of these experiments in desulphurization various -facts were established: - -(1) That sulphide of lead—especially in a pulverulent state—could not -be desulphurized in the same way as other sulphides, such as sulphides -of iron, copper, zinc, etc., because if roasted in a mechanical -furnace the temperature had to be kept low enough to avoid premature -sintering, which would choke the stirrers and cause trouble by the -ore clogging on the sides and bottom of the furnace. If, however, the -ore was roasted in a “dry state” at low temperature, a great deal of -sulphur remained in the product as sulphate of lead, which was as -bad for the subsequent blast-furnace work as the sulphide of lead -itself. When air was pressed through molten galena—in the same way as -through molten copper matte—a very heavy volatilization of lead took -place, while portions of it were reduced to metal or were contained as -sulphide in the molten matte, so that a good product was not obtained. - -(2) That no complete dead roast of lead ores could be obtained unless -the final product was thoroughly smelted and agglomerated. - -(3) That a well roasted lead ore could be obtained by oxidizing the PbS -with compressed air, after the ore had been suitably prepared. - -(4) That metal losses were mainly due to the excessive heat produced in -the oxidation of PbS to PbO, and other sulphides present in the ore. - -It was by making use of these facts that the H.-H. roasting process -was finally evolved, and by carefully applying its principles it is -possible to desulphurize completely the ore to a practically dead roast -of under 1 per cent. sulphur; in practice, however, such perfection -is unnecessary and a well agglomerated product with from 2 to 2.5 per -cent. sulphur is all that is required. During some trials in Australia, -where a great degree of perfection was aimed at, a block of over 2000 -tons of agglomerated, roasted ore was produced containing 1 per cent. -sulphur (as sulphide); as the ores contained an average of about 10 per -cent. Zn, this was a very fine result from a desulphurization point -of view, but it was not found that this 1 per cent. product gave any -better results in the subsequent smelting in the blast furnace than -later on a less carefully prepared material containing 2.5 per cent. -sulphur. - -In the early stages of experiment the great difficulty was to obtain -agglomeration without first fusing the sulphides in the ore, and -turning out a half-roasted product full of leady matte. Simple as the -thing now is, it seemed at times impossible to avoid this defect, and -it was only by a careful study of the effects of an addition of lime, -Fe₂O₃ or Mn₂O₃, and their properties that the right path -was struck. Before the introduction of the H.-H. process lime was -only used in the reverberatory process (Flintshire and Tarnowitz) to -stiffen the charge, but as Percy tells us that after its addition the -charge was glowing, it must have had a chemical as well as a mechanical -effect. In recognition of this fact fine caustic lime or crushed -limestone was mixed with the ore _before_ charging it into the furnace -and exposing it to an oxidizing heat. - -It was thought probable that a dioxide of lime might be temporarily -formed, which in contact with PbS would be decomposed immediately after -its formation, or that the CaO served as _Contactsubstanz_ in the same -way as spongy platinum, metallic silver, or oxide of iron. As CaSO₄ -and not CaSO₃ is always found in the roasted ore, this may prove -that CaO is really a contact substance for oxygen (see W. M. Hutchings, -_Engineering and Mining Journal_, Oct. 21, 1905, Vol. LXXX, p. 726). -The fact that the process works equally well with Fe₂O₃ instead -of CaO speaks against the theory of plumbate of lime. Whatever theory -may be correct, the fact remains that CaO assists the roasting process -and that by its use the premature agglomeration of the sulphide ore is -avoided. A further advantage of lime is that it keeps the charge more -porous and thus facilitates the passage of the air. - -The shape and size of the blowing apparatus best adapted for the -purpose in view occupied many months; starting from very shallow -pans or rectangular boxes several feet square with a few inches of -material over a perforated plate, it gradually resolved itself into the -cone-shaped receptacle—holding about a ton of ore—as first introduced -together with the process. In later years and in treating larger -quantities a more hemispherical form has been adopted, containing up to -15 tons of ore. - -It is probable about eight years were employed in actually working out -the process before it was introduced on any large scale at Pertusola, -but by the end of 1898 the greater part of the Pertusola ores were -treated by the process. Its first introduction to any other works was -in 1900, when it was started outside its home for the first time at -Braubach (Germany). Since then its application has gradually extended, -proceeding from Europe to Australia and Mexico and finally to America -and Canada, where recognition of its merits was more tardy than -elsewhere. It is now practically in general use all over the world and -is recognized as a sound addition to metallurgical progress. It is -doubtless only a step in the right direction and with its general use -a better knowledge of its principles will prevail, so that its future -development in one direction or another, as compared with present -results, may show some further progress. - -The present working of the H.-H. process still follows practically the -original lines laid down, and by preliminary roasting in a furnace -with lime, oxide of iron, or manganese (if not already contained in -the ore), prepares the ore for blowing in the converter. Mechanical -furnaces have been introduced to the entire exclusion of the old -hand-roasters, and the size of the converters has been gradually -increased from the original one-ton apparatus successively to 5, 7 -and 10 ton converters; at present some for 15 tons are being built in -Germany and will doubtless lead to a further economy. - -The mechanical furnace at present most in use is a single-hearth -revolving furnace with fixed rabbles, the latest being built with a -diameter of 26½ ft. and a relatively high arch to ensure a clear flame -and rapid oxidation of the ore. The capacity of these furnaces varies, -of course, with the nature of the ores to be treated, but with ordinary -lead ores (European and Australian practice) of from 50 per cent. to 60 -per cent. lead and 14 per cent, to 18 per cent. sulphur, the average -capacity may be taken at about 50 to 60 tons of crude ore per day of -24 hours. The consumption of coal with a well-constructed furnace is -very low and is always under 8 per cent.—6 per cent. being perhaps the -average. These furnaces require very little attention, being automatic -in their charging and discharging arrangements. - -The ore on leaving the furnace is charged into the converters by -various mechanical means (Jacob’s ladders, conveyors, etc.). The -converter charge usually consists of some hot ore direct from the -furnace, on top of which ore is placed which has been cooled down by -storage in bins or by the addition of water. The converter is generally -filled in two charges of five tons each, and the blowing time should -not be more than 4 to 6 hours. The product obtained should be porous -and well agglomerated, but easily broken up, tough melted material -being due to an excess of silica and too much lead sulphide. Attention, -therefore, to these two points (good preliminary roasting and -correction of the charge by lime) obviates this trouble. This roasted -ore should not contain more than about 1.5 to 2 per cent. sulphur, -and in a modern blast furnace gives surprisingly good results, the -matte-fall being in most cases reduced to nothing, and the capacity of -the furnace is largely increased, while the slags are poorer. - -If the converter charge has been properly prepared, the blowing -operation proceeds with the greatest smoothness and requires very -little attention on the part of the workmen, the heat and oxidation -rise gradually from the bottom and volatilization losses remain low, so -that it is possible, if desired, to produce hot concentrated sulphurous -gases suitable for the manufacture of sulphuric acid. - -Besides the actual economy obtained in roasting ores by the process, -a great feature of its success has been the remarkable improvement -in smelting and reducing the roasted ore as compared with previous -experience. This is due to the nature of the roasted material, which, -besides being much poorer in sulphur than was formerly the case, is -thoroughly porous and well agglomerated and contains—if the original -mixture is properly made—all the necessary slagging materials itself, -so that it practically becomes a case of smelting slags instead of ore, -and to an expert the difference is evident. - -Experience has shown that on an average the improvement in the capacity -of the blast furnace may be taken at about 50 to 100 per cent., so that -in works using the H.-H. process—after its complete introduction—about -half the blast furnaces formerly necessary for the same tonnage were -blown out. The matte-fall has become a thing of the past, so that, -except in those cases where some matte is required to collect the -copper contained in the ores, lead matte has disappeared and the -quantity of flue dust as well as the lead and silver losses have been -greatly reduced. - -Referring to the latest history of the H.-H. process, and the theory -of direct blowing, it may be remarked—putting aside all legal -questions—that the idea, metallurgically speaking, is attractive, as it -would seem that by eliminating one-half of the process and blowing the -ores direct without the expense of a preliminary roast a considerable -economy should be effected. Upon examination, however, this supposed -economy and simplicity is not at all of such great importance, and -in many cases, without doubt, would be retrogressive in lead ore -smelting rather than progressive. When costs of roasting in a furnace -are reduced to such a low figure as can be obtained by using 50 ton -furnaces and 10 to 15 ton converters, there is very little margin -for improvement in this direction. From the published accounts of -the Tarnowitz smelting works (the _Engineering and Mining Journal_, -Sept. 23, 1905, Vol. LXXX, p. 535) the cost of mechanical preliminary -roasting cannot exceed 25c. per ton, so that even assuming direct -blowing were as cheap as blowing a properly prepared material, the -total economy would only be the above figure, viz., 25c.; but this is -far from being the case. - -Direct blowing of a crude ore is considerably more expensive than -dealing with the H.-H. product, because of necessity the blowing -operation must be carried out slowly and with great care so as to avoid -heavy metal losses, and whereas a pre-roasted ore can be easily blown -in four hours and one man can attend to two or three 10 ton converters, -the direct blowing operation takes from 12 to 18 hours and requires the -continual attention of one man. In the first case the cost of labor -would be: One man at say $3 for 50 tons (at least), i.e., 6c. per -ton, and in the second case one man at $3 for 10 tons (at the best), -i.e., 30c., a difference in favor of pre-roasting of 24c., so that any -possible economy would disappear. Furthermore, as the danger of blowing -upon crude sulphides for 12 or 18 hours is greater as regards metal -losses than a quick operation of four hours, it is very likely that -instead of an economy there would be an increase in cost, owing to a -greater volatilization of metals. - -These remarks refer to ordinary lead ores with say 50 per cent. lead -and about 14 per cent. sulphur. With ores, however, such as are -generally treated in the United States the advantages of pre-roasting -are still more evident. These ores contain about 10 to 15 per cent. -lead, 30 to 40 per cent. sulphur, 20 to 30 per cent. iron, 10 per cent. -zinc, 5 per cent. silica, and lose the greater part of the pyritic -sulphur in the preliminary roasting, leaving the iron in the form of -oxide, which in the subsequent blowing operation acts in the same -way as lime. For this reason the addition of extra fluxes, such as -limestone, gypsum, etc., to the original ore is not necessary and only -a useless expense. - -In certain exceptional cases and with ores poor in sulphur, direct -blowing might be applicable, but for the general run of lead ores no -economy can be expected by doing away with the preliminary roast. - - - - - MAKING SULPHURIC ACID AT BROKEN HILL - - (August 11, 1904) - - -The Broken Hill Proprietary Company has entered upon the -manufacture of sulphuric acid on a commercial scale. The acid is -practically a by-product, being made from the gases emanating -from the desulphurization of the ores, concentrates, etc., by the -Carmichael-Bradford process. The acid can be made at a minimum of -cost, and thus materially enhances the value of the process recently -introduced for the separation of zinc blende from the tailings by -flotation. The following particulars are taken from a recently -published description of the process: The ores, concentrates, slimes, -etc., as the case may be, are mixed with gypsum, the quantity of the -latter varying from 15 to 25 per cent. The mixture is then granulated -to the size of marbles and dumped into a converter. The bottom of -the charge is heated from 400 to 500 deg. C. It is then subjected to -an induced current of air, and the auxiliary heat is turned off. The -desulphurization proceeds very rapidly with the evolution of heat and -the gases containing sulphurous anhydride. The desulphurization is very -thorough, and no losses occur through volatilization. The sulphur thus -rendered available for acid making is rather more than is contained in -the ore, the sulphur in the agglomerated product being somewhat less -than that accounted for by the sulphur contained in the added gypsum. -Thus from one ton of 14 per cent. sulphide ore it is possible to make -about 12 cwt. of chamber acid, fully equaling 7 cwt. of strong acid. - -The plant at present in use, which comprises a lead chamber of 40,000 -cu. ft., can turn out 35 tons of chamber acid per week. This plant is -being duplicated, and it has also been decided to erect a large plant -at Port Pirie for use in the manufacture of superphosphates. It is -claimed that the production of sulphuric acid from ores containing only -14 per cent. of sulphur establishes a new record. - - - - - THE CARMICHAEL-BRADFORD PROCESS - - BY DONALD CLARK - - (November 3, 1904) - - -Subsequent to the introduction of the Huntington-Heberlein process -in Australia, Messrs. Carmichael and Bradford, two employees of the -Broken Hill Proprietary Company, patented a process which bears their -name. Instead of starting with lime, or limestone and galena, as in -the Huntington-Heberlein process, they discovered that if sulphate of -lime is mixed with galena and the temperature raised, on blowing a -current of air through the mixture the temperature rises and the mass -is desulphurized. The process would thus appear to be a corollary of -the original one, and the reactions in the converter are identical. -Owing to the success of the acid processes in separating zinc sulphide -from the tailing at Broken Hill, it became necessary to manufacture -sulphuric acid locally in large quantity. The Carmichael-Bradford -process has been started for the purpose of generating the sulphur -dioxide necessary, and is of much interest as showing how gases rich -enough in SO₂ may be produced from a mixture containing only from 13 -to 16 per cent. sulphur. - -Gypsum is obtained in a friable state within about five miles from -Broken Hill. This is dehydrated, the CaSO, 2H₂O being converted into -CaSO₄ on heating to about 200 deg. C. The powdered residue is mixed -with slime produced in the milling operations and concentrate in the -proportion of slime 3 parts, concentrate 1 part, and lime sulphate 1 -part. The proportions may vary to some extent, but the sulphur contents -run from 13 to 16 or 17 per cent. The average composition of the -ingredients is as given in the table on the next page. - -These materials are moistened with water and well mixed by passing -them through a pug-mill. The small amount of water used serves to -set the product, the lime sulphate partly becoming plaster of paris, -2CaSO, H₂O. While still moist the mixture is broken into pieces not -exceeding two inches in diameter and spread out on a drying floor, -where excess of moisture is evaporated by the conjoint action of sun -and wind. - - ─────────────────┬─────┬───────────┬────────┬──────── - │SLIME│CONCENTRATE│CALCIUM │AVERAGE - │ │ │SULPHATE│ - ──────────────────┼─────┼───────────┼────────┼──────── - Galena │ 24 │ 70 │ │ 29 - Blende │ 30 │ 15 │ │ 21 - Pyrite │ 3 │ │ │ 2 - Ferric oxide │ 4 │ │ │ 2.5 - Ferrous oxide │ 1 │ │ │ 1 - Manganous oxide │ 6.5│ │ │ 5 - Alumina │ 5.5│ │ │ 3 - Lime │ 3.5│ │ 41 │ 10 - Silica │ 23 │ │ │ 14 - Sulphur trioxide │ │ │ 59 │ 12 - ─────────────────┴─────┴───────────┴────────┴──────── - -The pots used are small conical cast-iron ones, hung on trunnions, -and of the same pattern as used in the Huntington-Heberlein process. -Three of these are set in line, and two are at work while the third is -being filled. These pots have the same form of conical cover leading -to a telescopic tube, and all are connected to the same horizontal -pipe leading to the niter pots. Dampers are provided in each case. A -small amount of coal or fuel is fed into the pots and ignited by a -gentle blast; as soon as a temperature of about 400 to 500 deg. C. is -attained the dried mixture is fed in, until the pot is full; the cover -is closed down and the mass warms up. Water is first driven off, but -after a short time concentrated fumes of sulphur dioxide are evolved. -The amount of this gas may be as much as 14 per cent., but it is -usually kept at about 10 per cent., so as to have enough oxygen for -the conversion of the dioxide to the trioxide. The gases are led over -a couple of niter pots and thence to the usual type of lead chamber -having a capacity of 40,000 cu. ft. Chamber acid alone is made, since -this requires to be diluted for what is known as the saltcake process. - -The plant has now been in operation for some time and, it is claimed, -with highly successful results. The product tipped out of. the -converter is similar to that obtained in the Huntington-Heberlein -process, and is at once fit for the smelters, the amount of sulphur -left in it being always less than that originally introduced with the -gypsum; analysis of the desulphurized material shows usually from 3 to -4 per cent. sulphur. - - - - - THE CARMICHAEL-BRADFORD PROCESS - - BY WALTER RENTON INGALLS - - (October 28, 1905) - - -As described in United States patent No. 705,904, issued July 29, 1902, -lead sulphide ore is mixed with 10 to 35 per cent. of calcium sulphate, -the percentage varying according to the grade of the ore. The mixture -is charged into a converter and gradually heated externally until the -lower portion of the charge, say one-third to one-fourth, is raised to -a dull-red heat; or the reactions may be started by throwing into the -empty converter a shovelful of glowing coal and turning on a blast of -air sufficient to keep the coal burning and then feeding the charge -on top of the coal. This heating effects a reaction whereby the lead -sulphide of the ore is oxidized to sulphate and the calcium sulphate is -reduced to sulphide. The heated mixture being continuously subjected -to the blast of air, the calcium sulphide is re-oxidized to sulphate -and is thus regenerated for further use. This reaction is exothermic, -and sufficient heat is developed to complete the desulphurization of -the charge of ore by the concurrent reactions between the lead sulphate -(produced by the calcium sulphate) and portions of undecomposed ore, -sulphurous anhydride being thus evolved. The various reactions, which -are complicated in their nature, continue until the temperature of -the charge reaches a maximum, by which time the charge has shrunk -considerably in volume and has a tendency to become pasty. This becomes -more marked as the production of lead oxide increases, and as the -desired point of desulphurization is attained the mixture fuses; at -this stage the calcium sulphide which is produced from the sulphate -cannot readily oxidize, owing to the difficulty of coming into actual -contact with the air in the pasty mass, but, being subjected to the -strong oxidizing effect of the metallic oxide, it is converted into -calcium plumbate, while sulphurous anhydride is set free. The mass then -cools, as the exothermic reactions cease, and can be readily removed to -a blast furnace for smelting. - -The reactions above described are as outlined in the original -American patent specification. Irrespective of their accuracy, -the Carmichael-Bradford process is obviously quite similar to the -Huntington-Heberlein, and doubtless owes its origin to the latter. The -difference between them is that in the Huntington-Heberlein process -the ore is first partially roasted with addition of lime, and is then -converted in a special vessel. In the Carmichael-Bradford process -the ore is mixed with gypsum and is then converted directly. The -greatest claim for originality in the Carmichael-Bradford process -may be considered to lie in it as a method of desulphurizing gypsum, -inasmuch as not only is the sulphur of the ore expelled, but also a -part of the sulphur of the gypsum; and the sulphur is driven off as a -gas of sufficiently high tenor of sulphur dioxide to enable sulphuric -acid to be made from it economically. Up to the present time the -Carmichael-Bradford process has been put into practical use only at -Broken Hill, N. S. W. - -The Broken Hill Proprietary Company first conducted a series of tests -in a converter capable of treating a charge of 20 cwt. These tests were -made at the smelting works at Port Pirie. Exhaustive experiments made -on various classes of ores satisfactorily proved the general efficacy -of the process. The following ores were tried in these preliminary -experiments, viz.: - -First-grade concentrate containing: Pb, 60 per cent.; Zn, 10 per cent.; -S, 16 per cent.; Ag, 30 oz. - -Second-grade concentrate containing: Pb, 45 per cent.; Zn, 12.5 per -cent.; S, 14.5 per cent.; Ag, 22 oz. - -Slime containing: Pb, 21 per cent.; Zn, 17 per cent.; S, 13 per cent.; -Ag, 18 oz. - -Lead-copper matte containing: Fe, 42 per cent.; Pb, 17 per cent.; Zn, -13.3 per cent.; Cu, 2.4 per cent.; S, 23 per cent.; Ag, 25 oz. - -Other mattes, of varying composition up to 45 per cent. Pb and 100 oz. -Ag, were also tried. - -The results from these preliminary tests were so gratifying that a -further set of tests was made on lead-zinc slime, with a view of -ascertaining whether any volatilization losses occurred during the -desulphurization. This particular material was chosen because of its -accumulation in large proportions at the mine, and the unsatisfactory -result of the heap roasting which has recently been practised. The -heap roasting, although affording a product containing only 7 per cent. -S, which is delivered in lump form and therefore quite suitable for -smelting, resulted in a high loss of metal by volatilization (17 per -cent. Pb, 5 per cent. Ag). - -The result of nine charges of the slime treated by the -Carmichael-Bradford process was as follows: - - ─────────────────┬──────┬─────────────────────┬─────────────────────── - │ │ ASSAYS │ CONTENTS - │ Cwt. ├────┬──────┬────┬────┼─────┬─────┬────┬────── - │ │Pb% │Ag oz.│Zn% │ S% │ Pb │ Ag. │ Zn │ S - │ │ │ │ │ │cwt. │ oz. │cwt.│cwt. - ├──────┼────┼──────┼────┼────┼─────┼─────┼────┼────── - Raw slime │128.1 │21.3│ 18.0 │16.8│13.1│27.28│115.3│26.2│16.78 - Raw gypsum │ 54.9 │ │ │ │ │ │ │ │ 9.88 - ├──────┤ │ │ │ ├─────┼─────┼────┼────── - Total │183.0 │ │ │ │ │27.28│115.3│25.2│26.66 - ──────────────────┼──────┼────┼──────┼────┼────┼─────┼─────┼────┼────── - Sintered material│109.88│20.7│ 17.2 │ │4.80│22.74│ 94.5│ │ 5.27 - Middling │ 14.47│17.7│ 15.7 │ │6.20│ 2.56│ 11.3│ │ 0.89 - Fines │ 11.12│19.0│ 14.8 │ │7.50│ 2.11│ 8.2│ │ 0.83 - ├──────┼────┼──────┼────┼────┼─────┼─────┼────┼────── - Total │135.47│ │ │ │5.17│27.41│113.0│ │ 6.99 - ─────────────────┴──────┴────┴──────┴────┴────┴─────┴─────┴────┴────── - -These results indicated practically no volatilization of lead and -silver during the treatment, the lead showing a slight increase, viz., -0.47 per cent., and the silver 1.13 per cent. loss. A desulphurization -of 70.4 per cent. was effected. A higher desulphurization could have -been effected had this been desired. In the above tabulated results, -the term “middling” is applied to the loose fritted lumps lying on the -top of the charge: these are suitable for smelting, the fines being the -only portion which has to be returned. - -In order to test the practicability of making sulphuric acid, a plant -consisting of three large converters of capacity of five tons each, -together with a lead chamber 100 ft. by 20 ft. by 20 ft., was then -erected at Broken Hill, together with a dehydrating furnace, pug-mill, -and granulator. These converters are shown in the accompanying -engravings. - -A trial run was made with 108 tons of concentrate of the following -composition: 54 per cent. lead; 1.9 per cent. iron; 0.9 per cent. -manganese; 9.4 per cent. zinc; 14.6 per cent. sulphur; 19.2 per cent. -insoluble residue, and 24 oz. silver per ton. - -The converter charge consisted of 100 parts of the concentrate and -25 parts of raw gypsum, crushed to pass a 1 in. hole and retained -by a 0.25 in. hole, the material finer than 0.25 in. (which amounted -to 5 per cent. of the total) being returned to the pug-mill. After -desulphurization in the converter, the product assayed as follows: -48.9 per cent. lead; 1.80 per cent. iron; 0.80 per cent. manganese; -7.87 per cent. zinc; 3.90 per cent. sulphur; 1.02 per cent. alumina; -5.80 per cent. lime; 21.75 per cent. insoluble residue; 8.16 per cent. -undetermined (oxygen as oxides, sulphates, etc.); total, 100 per cent. -Its silver content was 22 oz. per ton. The desulphurized ore weighed -10 per cent. more than the raw concentrate. During this run 34 tons of -acid were made. - -A trial was then made on 75 tons of slime of the following composition: -18.0 per cent. lead; 16.6 per cent. zinc; 6.0 per cent. iron; 2.5 per -cent. manganese; 3.2 per cent. alumina; 2.1 per cent. lime; 38.5 per -cent. insoluble residue; total, 100 per cent. Its silver content was -19.2 oz. per ton. - -The converter charge in this case consisted of 100 parts of raw slime -and 30 parts of gypsum. The converted material assayed as follows: -16.1 per cent. lead; 14.0 per cent. zinc; 3.6 per cent. sulphur; 5.42 -per cent. iron; 2.25 per cent. manganese; 4.10 per cent. alumina; 8.60 -per cent. lime; 39.80 per cent. insoluble residue; 6.13 per cent. -undetermined (oxygen, etc.); total, 100 per cent.; and silver, 17.5 -oz. per ton. The increase in weight of desulphurized ore over that -of the raw ore was 11 per cent. During this run 22 tons of acid were -manufactured. - -The analysis of the gypsum used in each of the above tests (at Broken -Hill) was as follows: 76.1 per cent. CaSO₄, 2H₂O; 0.5 per cent. -Fe₂O₃; 4.5 per cent. Al₂O₃; 18.9 per cent. insoluble -residue. - -The plant was then put into continuous operation on a mixture of three -parts slime and one of concentrate, desulphurizing down to 4 per cent. -S, and supplying 20 tons of acid per week, and additions were made to -the plant as soon as possible. The acid made at Broken Hill has been -used in connection with the Delprat process for the concentration of -the zinc tailing. At Port Pirie, works are being erected with capacity -for desulphurization of about 35,000 tons per annum, with an acid -output of 10,000 tons. This acid is to be utilized for the acidulation -of phosphate rock. - -[Illustration: FIG. 15.—Details of Converters.] - -The cost of desulphurization of a ton of galena concentrate by the -Carmichael-Bradford process, based on labor at $1.80 per 8 hours, -gypsum at $2.40 per 2240 lb., and coal at $8.40 per 2240 lb., is -estimated as follows: - - 0.25 ton of gypsum $0.60 - Dehydrating and granulating gypsum .48 - Drying mixture of ore and gypsum .12 - Converting .24 - Spalling sintered material .12 - 0.01 ton coal .08 - ——-——- - Total $1.64 - -The lime in the sintered product is credited at 12c., making the net -cost $1.52 per ton (2240 lb.) of ore. - -The plant required for the Carmichael-Bradford process can be described -with sufficient clearness without drawings, except the converters. The -ore (concentrate, slime, etc.) to be desulphurized is delivered at the -top of the mill by cars, conveyors, or other convenient means, and -dumped into a bin. Two screw feeders placed inside the bin supply the -mill with ore, uniformly and as fast as it is required. These feeders -deliver the ore into a chute, which directs it into a vertical dry -mixer. - -A small bin, on the same level as the ore-bin, receives the crude -gypsum from cars. Thence it is fed automatically to a disintegrator, -which pulverizes it finely and delivers it into a storage bin -underneath. This disintegrator revolves at about 1700 r.p.m. and -requires 10 h.p. The body of the machine is cast iron, fitted with -renewable wearing plates (made of hard iron) in the grinding chamber. -The revolving parts consist of a malleable iron disc in which are fixed -steel beaters, faced on the grinding surface with highly tempered -steel. The bin that receives the floured gypsum contains a screw -conveyor similar to those in the ore-bin, and dumps the material into -push conveyors passing into the dehydrating furnace. They carry the -crushed gypsum along at a speed of about 1 ft. per minute, and allow -about 20 ft. to dehydrate the gypsum. This speed can, of course, be -regulated to suit requirements. - -The dehydrated gypsum runs down a chute into an elevator boot, and is -elevated into a bin which is on the same level as the ore-bin. This bin -also contains a screw conveyor, like that in the ore-bin. The speed of -delivery is regulated to deliver the right proportion of dehydrated -gypsum to the mixer. - -The mixer is of the vertical pattern and receives the sulphide ore -and dehydrated gypsum from the screw feeders. In it are set two flat -revolving cones running at different speeds, thus ensuring a thorough -mixture of the gypsum and ore. The mixed material drops from the -cones upon two baffle plates, and is wetted just before entering the -pug-mill. The pug-mill is a wrought-iron cylinder of ¼ in. plate about -2 ft. 6 in. diameter and 6 or 8 ft. long, and has the mixer fitted -to the head. It contains a 3 ft. wrought-iron spiral with propelling -blades, which forces the plastic mixture through ¾ in. holes in the -cover. The material comes out in long cylindrical pieces, but is broken -up and formed into marble-shaped pieces on dropping into a revolving -trommel. - -The trommel is about 5 ft. long, 2 ft. in diameter at the small end and -about 4 ft. at the large end. It revolves about a wrought-iron spindle -(2½ in. diameter) carrying two cast-iron hubs to which are fitted -arms for carrying the conical plate ⅛ in. thick. About 18 in. of -the small end of the cone is fitted with wire gauze, so as to prevent -the material as it comes out of the pug-mill from sticking to it. The -trommel is driven by bevel gearing at 20 to 25 r.p.m. The granulated -material formed in the trommel is delivered upon a drying conveyor. - -The conveyor consists of hinged wrought-iron plates flanged at the side -to keep the material from running off. It is driven from the head by -gearing, at a speed of 1 ft. per minute, passing through a furnace 10 -ft. long to dry and set the granules of ore and gypsum. This speed can, -of course, be regulated to suit requirements. The granulated material, -after leaving the furnace, is delivered to a single-chain elevator, -traveling at a speed of about 150 ft. per minute. It drops the material -into a grasshopper conveyor, driven by an eccentric, which distributes -the material over the length of a storage bin. From this bin the -material is directed into the converters by means of the chutes, which -have their bottom ends hinged so as to allow for the raising of the -hood when charging the converters. - -The converters are shown in the accompanying engravings, but they may -be of slightly different form from what is shown therein, i.e., they -may be more spherical than conical. They will have a capacity of about -four tons, being 6 ft. in diameter at the top, 4 ft. in diameter at -the false bottom, and about 5 ft. deep. They are swung on cast-iron -trunnions bolted to the body and turned by means of a hand-wheel and -worm (not shown). They are carried on strong cast-iron standards fitted -with bearings for trunnions, and all necessary brackets for tilting -gear. The hood has a telescopic funnel which allows it to be raised -or lowered, weights being used to balance it. At the apex of the cone -a damper is provided to regulate the draft. A 4 in. hole in the pot -allows the air from the blast-pipe, 18 in. in diameter, to enter under -the false perforated bottom, the connection between the two being made -by a flexible pipe and coupling. Two Baker blowers supply the blast for -the converters. The material, after being sintered, is tipped on the -floor in front of the converters and is there broken up to any suitable -size, and thence dispatched to the smelters. - -[Illustration: FIG. 16.—Arrangement of Converters.] - -The necessary power for a plant with a capacity of 150 tons of ore per -day will be supplied by a 50 h.p. engine. - - - - - THE SAVELSBERG PROCESS - - BY WALTER RENTON INGALLS - - (December 9, 1905) - - -There are in use at the present time three processes for the -desulphurization of galena by the new method, which has been referred -to as the “lime-roasting of galena.” The Huntington-Heberlein and the -Carmichael-Bradford processes have been previously described. The third -process of this type, which in some respects is more remarkable than -either of the others, is the invention of Adolf Savelsberg, director -of the smeltery at Ramsbeck, Westphalia, Germany, which is owned by -the Akt. Gesell. f. Bergbau, Blei. u. Zinkhüttenbetrieb zu Stolberg -u. in Westphalen. The process is in use at the Ramsbeck and Stolberg -lead smelteries of that company. It is described in American patent -No. 755,598, issued March 22, 1904 (application filed Dec. 18, 1903). -The process is well outlined in the words of the inventor in the -specification of that patent: - -“The desulphurizing of certain ores has been effected by blowing air -through the ore in a chamber, with the object of doing away with the -imperfect and costly process of roasting in ordinary furnaces; but -hitherto it has not been possible satisfactorily to desulphurize lead -ores in this manner, as, if air be blown through raw lead ores in -accordance with either of the processes used for treating copper ores, -for example, the temperature rises so rapidly that the unroasted lead -ore melts and the air can no longer act properly upon it, because -by reason of this melting the surface of the ores is considerably -decreased, the greater number of points or extent of surface which -the raw ore originally presented to the action of the oxygen of the -air blown through being lost, and, moreover, the further blowing -of air through the molten mass of ore produces metallic lead and a -plumbiferous slag (in which the lead oxide combines with the gangue) -and also a large amount of light dust, consisting mainly of sublimated -lead sulphide. Huntington and Heberlein have proposed to overcome -these objections by adopting a middle course, consisting in roasting -the ores with the addition of limestone for overcoming the ready -fusibility of the ores, and then subjecting them to the action of the -current of air in the chamber; but this process is not satisfactory, -because it still requires the costly previous operation in a roasting -furnace. - -[Illustration: Fig. 18.—Converter Ready to Dump.] - -“My invention is based on the observation which I have made that if -the lead ores to be desulphurized contain a sufficient quantity of -limestone it is possible, by observing certain precautions, to dispense -entirely with the previous roasting in a roasting furnace, and to -desulphurize the ores in one operation by blowing air through them. I -have found that the addition of limestone renders the roasting of the -lead ore unnecessary, because the limestone produces the following -effects: - -“The particles of limestone act mechanically by separating the -particles of lead ore from each other in such a way that premature -agglomeration is prevented and the whole mass is loosened and rendered -accessible to air; and, moreover, the limestone moderates the high -reaction temperature resulting from the burning of the sulphur, so -that the liquefaction of the galena, the sublimation of lead sulphide, -and the separation of metallic lead are avoided. The moderation of -the temperature of reaction is caused by the decomposition of the -limestone into caustic lime and carbon dioxide, whereby a large amount -of heat becomes latent. Further, the decomposition of the limestone -causes chemical reactions, lime being formed, which at the moment of -its formation is partly converted into sulphate of lime at the expense -of the sulphur contained in the ore, and this sulphate of lime, when -the scorification takes place, is transformed into calcium silicate -by the silicic acid, the sulphuric acid produced thereby escaping. -The limestone also largely contributes to the desulphurization of the -ore, as it causes the production of sulphuric acid at the expense of -the sulphur of the ore, which sulphuric acid is a powerful oxidizing -agent. If, therefore, a mixture of raw lead ore and limestone (which -mixture must, of course, contain a sufficient amount of silicic acid -for forming silicates) be introduced into a chamber and a current of -air be blown through the mixture, and at the same time the part of the -mixture which is near the blast inlet be ignited, the combustion of the -sulphur will give rise to very energetic reactions, and sulphurous -acid, sulphuric acid, lead oxide, sulphates and silicates are produced. -The sulphurous acid and the carbon dioxide escape, while the sulphuric -acid and sulphates act in their turn as oxidizing agents on the -undecomposed galena. Part of the sulphates is decomposed by the silicic -acid, thereby liberating sulphuric acid, which, as already stated, acts -as an oxidizing agent. The remaining lead oxide combines finally with -the gangue of the ore and the non-volatile constituents of the flux -(the limestone) to form the required slag. These several reactions -commence at the blast inlet at the bottom of the chamber, and extend -gradually toward the upper portion of the charge of ore and limestone. -Liquefaction of the ores does not take place, for although a slag is -formed it is at once solidified by the blowing in of the air, the -passages formed thereby in the hardening slag allowing of the continued -passage therethrough of the air. The final product is a silicate -consisting of lead oxide, lime, silicic acid, and other constituents of -the ore, which now contains but little or no sulphur and constitutes a -coherent solid mass, which, when broken into pieces, forms a material -suitable to be smelted. - -“The quantity of limestone required for the treatment of the lead -ores varies according to the constitution of the ores. It should, -however, amount generally to from 15 to 20 per cent. As lead ores do -not contain the necessary amount of limestone as a natural constituent, -a considerable amount of limestone must be added to them, and this -addition may be made either during the dressing of the ores or -subsequently. - -“For the satisfactory working of the process, the following precautions -are to be observed: In order that the blowing in of the air may not -cause particles of limestone to escape in the form of dust before -the reaction begins, it is necessary to add to the charge before it -is subjected to the action in the chamber a considerable amount of -water—say 5 per cent. or more. This water prevents the escape of dust, -and it also contributes considerably to the formation of sulphuric -acid, which, by its oxidizing action, promotes the reaction, and, -consequently, also the desulphurization. It is advisable, in conducting -the operation, not to fill the chamber with the charge at once, but -first only partly to fill it and add to the charge gradually while the -chamber is at work, as by this means the reaction will take place more -smoothly in the mass. - -[Illustration: Fig. 19.—Charge Dumped.] - -“It is advantageous to proceed as follows: The bottom part of a -chamber of any suitable form is provided with a grate, on which is -laid and ignited a mixture of fuel (coal, coke, or the like) and -pieces of limestone. By mixing the fuel with pieces of limestone the -heating power of the fuel is reduced and the grate is protected, -while at the same time premature melting of the lower part of the -charge is prevented; or the grate may be first covered with a layer -of limestone and the fuel be laid thereon, and then another layer of -limestone be placed on the fuel. On the material thus placed in the -chamber, a uniform charge of lead ore and limestone—say about 12 in. -high—is placed, this having been moistened as previously explained. -Under the influence of the air-blast and the heat, the reactions -hereinbefore described take place. When the upper surface of the first -layer becomes red-hot, a further charge is laid thereon, and further -charges are gradually introduced as the surface of the preceding -charge becomes red-hot, until the chamber is full. So long as charges -are still introduced a blast of air of but low pressure is blown -through; but when the chamber is filled a larger quantity of air at a -higher pressure is blown through. The scorification process then takes -place, a very powerful desulphurization having preceded it. During the -scorification the desulphurization is completed. - -“When the process is completed, the chamber is tilted and the -desulphurized mass falls out and is broken into small pieces for -smelting.” - -The drawing on page 190, Fig. 17, shows a side view of the apparatus -used in connection with the process, which will be readily understood -without special description. The dotted lines show the pot in its -emptying position. The series of operations is clearly illustrated in -Figs. 18-20, which are reproduced from photographs. - -This process has now been in practical use at Ramsbeck for three years, -where it is employed for the desulphurization of galena of high grade -in lead, with which are mixed quartzose silver ore (or sand if no such -ore be available), and calcareous and ferruginous fluxes. A typical -charge is 100 parts of lead ore, 10 parts of quartzose silver ore, -10 parts of spathic iron ore, and 19 parts of limestone. A thorough -mixture of the components is essential; after the mixture has been -effected, the charge is thoroughly wetted with about 5 per cent. -of water, which is conceived to play a threefold function in the -desulphurizing operation, namely: (1) preservation of the homogeneity -of the mixture during the blowing; (2) reduction of temperature during -the process; and (3) formation of sulphuric acid in the process, which -promotes the desulphurization of the ore. - -[Illustration: FIG. 17.—Savelsberg Converter.] - -The moistened charge is conveyed to the converters, into which it -is fed in thin layers. The converters are hemispherical cast-iron -pots, supported by trunnions on a truck, as shown in the accompanying -engravings. Except for this method of support, which renders the -pot movable, the arrangement is quite similar to that which is -employed in the Huntington-Heberlein process. The pots which are now -in use at Ramsbeck have capacity for about 8000 kg. of charge, but -it is the intention of the management to increase the capacity to -10,000 or 12,000 kg. Previously, pots of only 5000 kg. capacity were -employed. Such a pot weighed 1300 kg., exclusive of the truck. The -air-blast was about 7 cu. m. (247.2 cu. ft.) per min., beginning at -a pressure of 10 to 20 cm. of water (2¾ to 4½ oz.) and rising to -50 to 60 cm. (11½ to 13½ oz.) when the pot was completely filled with -charge. The desulphurization of a charge is completed in 18 hours. A -pot is attended by one man per shift of 12 hours; this is only the -attention of the pot proper, the labor of conveying material to it and -breaking up the desulphurized product being extra. One man per shift -should be able to attend to two pots, which is the practice in the -Huntington-Heberlein plants. - -[Illustration: Fig. 20.—Converter in Position for Blowing.] - -When the operation in the pot is completed, the latter is turned on its -trunnions, until the charge slides out by gravity, which it does as a -solid cake. This is caused to fall upon a vertical bar, which breaks -it into large pieces. By wedging and sledging these are reduced to -lumps of suitable size for the blast furnace. When the operation has -been properly conducted the charge is reduced to about 2 or 3 per cent. -sulphur. It is expected that the use of larger converters will show -even more favorable results in this particular. - -As in the Huntington-Heberlein and Carmichael-Bradford processes, one -of the greatest advantages of the Savelsberg process is the ability to -effect a technically high degree of desulphurization with only a slight -loss of lead and silver, which is of course due to the perfect control -of the temperature in the process. The precise loss of lead has not yet -been determined, but in the desulphurization of galena containing 60 -to 78 per cent. lead, the loss of lead is probably not more than 1 per -cent. There appears to be no loss of silver. - -The process is applicable to a wide variety of lead-sulphide ores. The -ore treated at Ramsbeck contains 60 to 78 per cent. lead and about -15 per cent. of sulphur, but ore from Broken Hill, New South Wales, -containing 10 per cent. of zinc has also been treated. A zinc content -up to 7 or 8 per cent. in the ore is no drawback, but ores carrying a -higher percentage of zinc require a larger addition of silica and about -5 per cent. of iron ore in order to increase the fusibility of the -charge. The charge ordinarily treated at Ramsbeck is made to contain -about 11 per cent. of silica. The presence of pyrites in the ore is -favorable to the desulphurization. Dolomite plays the same part in -the process that limestone does, but is of course less desirable, in -view of the subsequent smelting in the blast furnace. The ore is best -crushed to about 3 mm. size, but good results have been obtained with -ore coarser in size than that. However, the proper size is somewhat -dependent upon the character of the ore. The blast pressure required in -the converter is also, of course, somewhat dependent upon the porosity -of the charge. Fine slimes are worked up by mixture with coarser ore. - -In making up the charge, the proportion of limestone is not varied -much, but the proportions of silica and iron must be carefully modified -to suit the ore. Certain kinds of ore have a tendency to remain -pulverulent, or to retain balls of unsintered, powdered material. -In such cases it is necessary to provide more fusible material in -the charge, which is done by varying the proportions of silica and -iron. The charge must, moreover, be prepared in such a manner that -overheating, and consequently the troublesome fusion of raw galena, -will be avoided. - -The essential difference between the Huntington-Heberlein and -Savelsberg processes is the use in the former of a partially -desulphurized ore, containing lime and sulphate of lime; and the use -in the latter of raw ore and carbonate of lime. It is claimed that the -latter, which loses its carbon dioxide in the converter, necessarily -plays a different chemical part from that of quicklime or gypsum. -Irrespective of the reactions, however, the Savelsberg process has the -great economic advantage of dispensing with the preliminary roasting of -the Huntington-Heberlein process, wherefore it is cheaper both in first -cost of plant and in operation. - - - - - THE LIME-ROASTING OF GALENA[32] - - BY WALTER RENTON INGALLS - - -During the last two years, and especially during the last six -months, a number of important articles upon the new methods for the -desulphurization of galena have been published in the technical -periodicals, particularly in the _Engineering and Mining Journal_ -and in _Metallurgie_. I proposed for these methods the type-name -of “lime-roasting of galena,” as a convenient metallurgical -classification,[33] and this term has found some acceptance. The -articles referred to have shown the great practical importance of these -new processes, and the general recognition of their metallurgical and -commercial value, which has already been accorded to them. It is my -present purpose to review broadly the changes developed by them in -the metallurgy of lead, in which connection it is necessary to refer -briefly to the previous state of the art. - -The elimination of the sulphur content of galena has been always the -most troublesome part of the smelting process, being both costly in the -operation and wasteful of silver and lead. Previous to the introduction -of the Huntington-Heberlein process at Pertusola, Italy, it was -effected by a variety of methods. In the treatment of non-argentiferous -galena concentrate, the smelting was done by the roast-reduction method -(roasting in reverberatory furnace and smelting in blast furnace); -the roast-reaction method, applied in reverberatory furnaces; and the -roast-reaction method, applied in Scotch hearths.[34] Precipitation -smelting, simple, had practically gone out of use, although its -reactions enter into the modern blast-furnace practice, as do also -those of the roast-reaction method. - -In the treatment of argentiferous lead ores, a combination of the -roast-reduction, roast-reaction and precipitation methods had been -developed. Ores low in lead were still roasted, chiefly in hand-worked -reverberatories (the mechanical furnaces not having proved well adapted -to lead-bearing ores), while the high loss of lead and silver in -sinter-or slag-roasting of rich galenas had caused those processes to -be abandoned, and such ores were charged raw into the blast furnace, -the part of their sulphur which escaped oxidation therein reappearing -in the form of matte. In the roast-reduction smelting of galena alone, -however, there was no way of avoiding the roasting of the whole, or at -least a very large percentage of the ore, and in this roasting the ore -had necessarily to be slagged or sintered in order to eliminate the -sulphur to a satisfactory extent. This is exemplified in the treatment -of the galena concentrate of southeastern Missouri at the present time. - -Until the two new Scotch-hearth plants at Alton and Collinsville, Ill., -were put in operation, the three processes of smelting the southeastern -Missouri galena were about on an equal footing. Their results per ton -of ore containing 65 per cent. lead were approximately as follows[35]: - - ──────────────────┬──────────────┬──────────── - METHOD │ COST │ EXTRACTION - ──────────────────┼──────────────┼─────────── - Reverberatory │ $6.50-7.00 │ 90-92% - Scotch hearth │ 5.75-6.50 │ 87-88% - Roast-reduction │ 6.00-7.00 │ 90-92% - ──────────────────┴──────────────┴─────────── - -The new works employ the Scotch-hearth process, with bag-houses for -the recovery of the fume, which previously was the weak point of this -method of smelting.[36] This improvement led to a large increase in the -recovery of lead, so that the entire extraction is now approximately 98 -per cent. of the content of the ore, while on the other hand the cost -of smelting per ton of ore has been reduced through the increased size -of these plants and the introduction of improved means for handling -ore and material. The practice of these works represents the highest -efficiency yet obtained in this country in the smelting of high-grade -galena concentrate, and probably it cannot be equaled even by the -Huntington-Heberlein and similar processes. The Scotch-hearth and -bag-house process is therefore the one of the older methods of smelting -which will survive. - -In the other methods of smelting, a large proportion of the cost is -involved in the roasting of the ore, which amounts in hand-worked -reverberatory furnaces to $2 to $2.50 per ton. Also, the larger -proportion of the loss of metal is suffered in the roasting of the ore, -this loss amounting to from 6 to 8 per cent. of the metal content of -such ore as is roasted. The loss of lead in the combined process of -treatment depends upon the details of the process. The chief advantage -of lime-roasting in the treatment of this class of ore is in the higher -extraction of metal which it affords. This should rise to 98 per cent. -That figure has been, indeed, surpassed in operations on a large scale, -extending over a considerable period. - -In the treatment of the argentiferous ores of the West different -conditions enter into the consideration. In the working of those ores, -the present practice is to roast only those which are low in lead, -and charge raw into the blast furnace the rich galenas. The cost of -roasting is about $2 to $2.50 per ton; the cost of smelting is about -$2.50 per ton. On the average about 0.4 ton of ore has to be roasted -for every ton that is smelted. The cost of roasting and smelting is -therefore about $3.50 per ton. In good practice the recovery of silver -is about 98 per cent. and of lead about 95 per cent., reckoned on basis -of fire assays. - -In treatment of these ores, the lime-roasting process offers several -advantages. It may be performed at less than the cost of ordinary -roasting.[37] The loss of silver and lead during the roasting is -reduced to insignificant proportion. The sulphide fines which must be -charged raw into the blast furnace are eliminated, inasmuch as they -can be efficiently desulphurized in the lime-roasting pots without -significant loss; all the ore to be smelted in the blast furnace -can be, therefore, delivered to it in lump form, whereby the speed -of the blast furnace is increased and the wind pressure required -is decreased. Finally, the percentage of sulphur in the charge is -reduced, producing a lower matte-fall, or no matte-fall whatever, with -consequent saving in expense of retreatment. In the case of a new -plant, the first cost of construction and the ground-space occupied -are materially reduced. Before discussing more fully the extent and -nature of these savings, it is advisable to point out the differences -among the three processes of lime-roasting that have already come into -practical use. - -In the Huntington-Heberlein process, the ore is mixed with suitable -proportions of limestone and silica (or quartzose ore) and is then -partially roasted, say to reduction of the sulphur to one half. The -roasting is done at a comparatively low temperature, and the loss of -metals is consequently small. The roasted ore is dampened and allowed -to cool. It is then charged into a hemispherical cast-iron pot, with -a movable hood which covers the top and conveys off the gases. There -is a perforated grate in the bottom of the pot, on which the ore -rests, and air is introduced through a pipe entering the bottom of the -pot, under the grate. A small quantity of red-hot calcines from the -roasting furnaces is thrown on the grate to start the reaction; a layer -of cold, semi-roasted ore is put upon it, the air blast is turned on -and reaction begins, which manifests itself by the copious evolution -of sulphur fumes. These consist chiefly of sulphur dioxide, but they -contain more or less trioxide, which is evident from the solution of -copperas that trickles from the hoods and iron smoke-pipes, wherein the -moisture condenses. As the reaction progresses, and the heat creeps -up, more ore is introduced, layer by layer, until the pot is full. -Care is taken by the operator to compel the air to pass evenly and -gently through the charge, wherefore he is watchful to close blow-holes -which develop in it. At the end of the operation, which may last from -four to eighteen hours, the ore becomes red-hot at the top. The hood -is then pushed up, and the pot is turned on its trunnions, by means -of a hand-operated wheel and worm-gear, until the charge slides out, -which it does as a solid, semi-fused cake. The pot is then turned back -into position. Its design is such that the air-pipe makes automatic -connection, a flanged pipe cast with the pot settling upon a similarly -flanged pipe communicating with the main, a suitable gasket serving -to make a tight joint. The pots are set at an elevation of about 12 -ft. above the ground, so that when the charge slides out the drop will -break it up to some extent, and it is moreover caused to fall on a -wedge, or similar contrivance, to assist the breakage. After cooling it -is further broken up to furnace size by wedging and sledging; the lumps -are forked out, and the fines screened and returned to a subsequent -charge for completion of their desulphurization. - -The Savelsberg process differs from the Huntington-Heberlein in respect -to the preliminary roasting, which in the Savelsberg process is -omitted, the raw ore, mixed with limestone and silica, being charged -directly into the converter. The Savelsberg converter is supported on -a truck, instead of being fixed in position, but otherwise its design -and management are quite similar to those of the Huntington-Heberlein -converter. In neither case are there any patents on the converters. -The patents are on the processes. In view of the litigation that -has already been commenced between their respective owners, it is -interesting to examine the claims. - -The Huntington-Heberlein patent (U. S. 600,347, issued March 8, 1898, -applied for Dec. 9, 1896) has the following claims: - -1. The herein-described method of oxidizing sulphide ores of lead -preparatory to reduction to metal, which consists in mixing with the -ore to be treated an oxide of an alkaline-earth metal, such as calcium -oxide, subjecting the mixture to heat in the presence of air, then -reducing the temperature and finally passing air through the mass -to complete the oxidation of the lead, substantially as and for the -purpose set forth. - -2. The herein-described method of oxidizing sulphide ores of lead -preparatory to reduction to metal, which consists in mixing calcium -oxide or other oxide of an alkaline-earth metal with the ore to be -treated, subjecting the mixture in the presence of air to a bright-red -heat (about 700 deg. C.), then cooling down the mixture to a dull-red -heat (about 500 deg. C.), and finally forcing air through the mass -until the lead ore, reduced to an oxide, fuses, substantially as set -forth. - -3. The herein-described method of oxidizing lead sulphide in the -preparation of the same for reduction to metal, which consists in -subjecting the sulphide to a high temperature in the presence of an -oxide of an alkaline-earth metal, such as calcium oxide, and oxygen, -and then lowering the temperature substantially as set forth. - -Adolf Savelsberg, in U. S. patent 755,598 (issued March 22, 1904, -applied for Dec. 18, 1903) claims: - -1. The herein-described process of desulphurizing lead ores, which -consists in mixing raw ore with limestone and then subjecting the -mixture to the simultaneous application of heat and a current of air in -sufficient proportions to substantially complete the desulphurization -in one operation, substantially as described. - -2. The herein-described process of desulphurizing lead ores, which -process consists in first mixing the ores with limestone, then -moistening the mixture, then filling it without previous roasting into -a chamber, then heating it and treating it by a current of air, as and -for the purpose described. - -3. The herein-described process of desulphurizing lead ores, which -consists in mixing raw ores with limestone, then filling the mixture -into a chamber, then subjecting the mixture to the simultaneous -application of heat and a current of air in sufficient proportions -to substantially complete the desulphurization in one operation, the -mixture being introduced into the chamber in partial charges introduced -successively at intervals during the process, substantially as -described. - -4. The herein-described process of desulphurizing lead ores, then -moistening the mixture, then filling it without previous roasting into -a chamber, then heating it and treating it by a current of air, the -mixture being introduced into the chamber in partial charges introduced -successively at intervals during the process, as and for the purpose -described. - -5. The herein-described process of desulphurizing lead ores, which -process consists in first mixing the ores with sufficient limestone to -keep the temperature of the mixture below the melting-point of the ore, -then filling the mixture into a chamber, then heating said mixture and -treating it with a current of air, as and for the purpose described. - -6. The herein-described process of desulphurizing lead ores, which -process consists in first mixing the ores with sufficient limestone to -mechanically separate the particles of galena sufficiently to prevent -fusion, and to keep the temperature below the melting-point of the ore -by the liberation of carbon dioxide, then filling the mixture into a -chamber, then heating said mixture and treating it with a current of -air, as and for the purpose described. - -The Carmichael-Bradford process differs from the Savelsberg by the -treatment of the raw ore mixed with gypsum instead of limestone, -and differs from the Huntington-Heberlein both in respect to the -use of gypsum and the omission of the preliminary roasting. The -Carmichael-Bradford process has not been threatened with litigation, -so far as I am aware. The claims of its original patent read as -follows[38]: - -1. The process of treating mixed sulphide ores, which consists in -mixing with said ores a sulphur compound of a metal of the alkaline -earths, starting the reaction by heating the same, thereby oxidizing -the sulphide and reducing the sulphur compound of the alkali metal, -passing a current of air to oxidize the reduced sulphide compound of -the metal of the alkalies preparatory to acting upon a new charge of -sulphide ores, substantially as and for the purpose set forth. - -2. The process of treating mixed sulphide ores, which consists in -mixing calcium sulphate with said ores, starting the reaction by -means of heat, thereby oxidizing the sulphide ores, liberating -sulphurous-acid gas and converting the calcium sulphate into calcium -sulphide and oxidizing the calcium sulphide to sulphate preparatory to -treating a fresh charge of sulphide ores, substantially as and for the -purpose set forth. - -The process described by W. S. Bayston, of Melbourne (Australian patent -No. 2862), appears to be identical with that of Savelsberg. - -Irrespective of the validity of the Savelsberg and Carmichael-Bradford -patents, and without attempting to minimize the ingenuity of their -inventors and the importance of their discoveries, it must be conceded -that the merit for the invention and introduction of lime-roasting of -galena belongs to Thomas Huntington and Ferdinand Heberlein. The former -is an American, and this is the only claim that the United States can -make to a share in this great improvement in the metallurgy of lead. It -is to be regretted, moreover, that of all the important lead-smelting -countries in the world, America has been the most backward in adopting -it. - -The details of the three processes and the general results accomplished -by them have been rather fully described in a series of articles -recently published in the _Engineering and Mining Journal_. There -has been, however, comparatively little discussion as to costs; and -unfortunately the data available for analysis are extremely scanty, due -to the secrecy with which the Huntington-Heberlein process, the most -extensively exploited of the three, has been veiled. Nevertheless, I -may attempt an approximate estimation of the various details, taking -the Huntington-Heberlein process as the basis. - -The ore, limestone and silica are crushed to pass a four-mesh screen. -This is about the size to which it would be necessary to crush as -preliminary to roasting in the ordinary way, wherefore the only -difference in cost is the charge for crushing the limestone and silica, -which in the aggregate may amount to one-sixth of the weight of the raw -sulphide and may consequently add 2 to 2.5c. to the cost of treating -a ton of ore. The mixing of ore and fluxes may be costly or cheap, -according to the way of doing it. If done in a rational way it ought -not to cost more than 10c. per ton of ore, and may come to less. The -delivery of the ore from the mixing-house to the roasting furnaces -ought to be done entirely by mechanical means, at insignificant cost. - -The Heberlein roasting furnace, which is used in connection with the -H.-H. process, is simply an improvement on the old Brunton calciner—a -circular furnace, with revolving hearth. The construction of this -furnace, according to American designs, is excellent. The hearth is -26 ft. in diameter; it is revolved at slow speed and requires about -1.5 h.p. A flange at the periphery of the hearth dips into sand in an -annular trough, thus shutting off air from the combustion chamber, -except through the ports designed for its admittance. The mechanical -construction of the furnace is workmanlike, and the mechanism under the -hearth is easy of access and comfortably attended to. - -A 26 ft. furnace roasts about 80,000 lb. of charge per 24 hours. In -dealing with an ore containing 20 to 22 per cent. of sulphur, the -latter is reduced to about 10 to 11 per cent., the consumption of -coal being about 22.5 per cent. of the weight of the charge. The -hearth efficiency is about 150 lb. per sq. ft., which in comparison -with ordinary roasting is high. The coal consumption, however, is not -correspondingly low. Two furnaces can be managed by one man per 8 hour -shift. On the basis of 80 tons of charge ore per 24 hours, the cost of -roasting should be approximately as follows: - - Labor—3 men at $2.50 $ 7.50 - Coal—18 tons at $2 36.00 - Power 3.35 - Repairs 3.35 - —————— - Total $50.20 = 63c. per ton. - -In the above estimate repairs have been reckoned at the same figure as -is experienced with Brückner cylinders, and the cost of power has been -allowed for with fair liberality. The estimated cost of 63c. per ton -is comparable with the $1.10 to $1.45 per ton, which is the result of -roasting in Brückner cylinders in Colorado, reducing the ore to 4.5-6 -per cent. sulphur. - -The Heberlein furnace is built up to considerable elevation above -the ground level, externally somewhat resembling the Pearce turret -furnace. This serves two purposes: (1) it affords ample room under the -hearth for attention to the driving mechanism; and (2) it enables the -ore to be discharged by gravity into suitable hoppers, without the -construction of subterranean gangways. The ore discharges continuously -from the furnace, at dull-red heat, into a brick bin, wherein it is -cooled by a water-spray. Periodically a little ore is diverted into a -side bin, in which it is kept hot for starting a subsequent charge in -the converter. - -The cooled ore is conveyed from the receiving bins at the roasting -furnaces to hopper-bins above the converters. If the tramming be done -by hand the cost, with labor at 25c. per hour, may be approximately -12.5c. per ton of ore, but this should be capable of considerable -reduction by mechanical conveyance. - -The converters are hemispherical pots of cast iron, 9 ft. in -diameter at the top, and about 4 ft. in depth. They are provided -with a circular, cast-iron grate, which is ¾ in. thick and 6 ft. in -diameter and is set and secured horizontally in the pot. This grate -is perforated with holes ¾ in. in diameter, 2 in. apart, center to -center, and is similar to the Wetherill grate employed in zinc oxide -manufacture. The pot itself is about 2½ in. thick at the bottom, -thinning to about 1½ in. at the rim. It is supported on trunnions and -is geared for convenient turning by hand. The blast pipe which enters -the pot at the bottom is 6 in. in diameter. - -Two roasting furnaces and six converters are rated nominally as a 90 -ton plant. This rating is, however, considerably in excess of the -actual capacity, at least on certain ores. The time required for -desulphurization in the converter apparently depends a good deal upon -the character of the ore. The six converters may be arranged in a -single row, or in two rows of three in each. They are set so that the -rim of the pot, when upright, is about 12 ft. above the ground level. -A platform gives access to the pots. One man per shift can attend to -two pots. His work consists in charging them, which is done by gravity, -spreading out the charge evenly in the pot, closing any blow-holes -which may develop, and at the end of the operation raising the hood -(which covers the pot during the operation) and dumping the pot. The -work is easy. The conditions under which it is done are comfortable, -both as to temperature and atmosphere. Reports have shown a great -reduction in liability to lead-poisoning in the works where the H.-H. -process has been introduced. - -A new charge is started by kindling a small wood or coal fire on the -grate, then throwing in a few shovelfuls of hot calcines, and finally -dropping in the regular charge of damp ore (plus the fluxes previously -referred to). The charge is introduced in stages, successive layers -being dropped in and spread out as the heat rises. At the beginning -the blast is very low—about 2 oz. It is increased as the hight of the -ore in the pot rises, finally attaining about 16 oz. The operation -goes on quietly, the smoke rising from the surface evenly and gently, -precisely as in a well-running blast furnace. While the charge is still -black on top, the hand can be held with perfect comfort, inside of -the hood, immediately over the ore. This explains, of course, why the -volatilization of silver and lead is insignificant. There is, moreover, -little or no loss of ore as dust, because the ore is introduced damp, -and the passage of the air through it is at low velocity. In the -interior of the charge, however, there is high temperature (evidently -much higher than has been stated in some descriptions), as will be -shown further on. The conditions in this respect appear to be analogous -to those of the blast furnace, which, though smelting at a temperature -of say 1200 deg. C. at the level of the tuyeres, suffers only a slight -loss of silver and lead by volatilization. - -At the end of the operation in the H.-H. pot, the charge is dull red -at the top, with blow-holes, around which the ore is bright red. -Imperfectly worked charges show masses of well-fused ore surrounded -by masses of only partially altered ore, a condition which may be -ascribed to the irregular penetration of air through the charge, -affording good evidence of the important part which air plays in the -process. A properly worked charge is tipped out of the pot as a solid -cake, which in falling to the ground breaks into a few large pieces. -As they break, it appears that the interior of the charge is bright -red all through, and there is a little molten slag which runs out of -cavities, presumably spots where the chemical action has been most -intense. When cold, the thoroughly desulphurized material has the -appearance of slag-roasted galena. Prills of metallic lead are visible -in it, indicating reaction between lead sulphide and lead sulphate. - -The columns of the structure supporting the pots should be of steel, -since fragments of the red-hot ore dumped on the ground are likely to -fall against them. To hasten the cooling of the ore, water is sometimes -played on it from a hose. This is bad, since some is likely to splash -into the still inverted pot, leading to cracks. The cracked pots at -certain works appear to be due chiefly to this cause, in the absence of -which the pots ought to last a long time, inasmuch as the conditions -to which they are subjected during the blowing process are not at all -severe. When the ore is sufficiently cold it is further broken up, -first by driving in wedges, and finally by sledging down to pieces -of orange size, or what is suitable for the blast furnace. These are -forked out, leaving the fine ore, which comes largely from the top of -the charge and is therefore only partially desulphurized. The fines -are, therefore, re-treated with a subsequent charge. The quantity is -not excessive; it may amount to 7 or 8 per cent. of the charge. - -The breaking up of the desulphurized ore is one of the problems of the -process, the necessity being the reduction of several large pieces -of fused, or semi-fused, material weighing two or three tons each. -When done by hand only, as is usually (perhaps always) the practice, -the operation is rather expensive. It would appear, however, to -be not a difficult matter to devise some mechanical aids for this -process—perhaps to make it entirely mechanical. When done by hand, a -6-pot plant requires 6 men per shift sledging and forking. With 8-hour -shifts, this is 18 men for the breaking of about 60 tons of material, -which is about 3⅓ tons per man per 8 hours. With labor at 25c. per -hour, the cost of breaking the fused material comes to 60c. per ton. It -may be remarked, for comparison, that in breaking ore as it ordinarily -comes, coarse and fine together, a good workman would normally be -expected to break 5 to 5.5 tons in a shift of 8 hours. - -The ordinary charge for the standard converter is about 8 tons (16,000 -lb.) of an ore weighing 166 lb. per cu. ft. With a heavier ore, like -a high-grade galena, the charge would weigh proportionately more. The -time of working off a charge is decidedly variable. Accounts of the -operation of the process in Australia tell of charge-workings in 3 -to 5 hours, but this does not correspond with the results reported -elsewhere, which specify times of 12 to 18 hours. Assuming an average -of 16 hours, which was the record of one plant, six converters would -have capacity for about 72 tons of charge per 24 hours, or about 58 -tons of ore, the ratio of ore to flux being 4:1. The loss in weight -of the charge corresponds substantially to the replacement of sulphur -by oxygen, and the expulsion of carbon dioxide. The finished charge -contains on the average from 3 to 5 per cent. sulphur. This is -about the same as the result achieved in good practice in roasting -lead-bearing ores in hand-worked reverberatory furnaces, but curiously -the H.-H. product, in some cases at least, does not yield any matte, -to speak of, in the blast furnace; the product delivered to the latter -being evidently in such condition that the remaining sulphur is almost -completely burned off in the blast furnace. This is an important saving -effected by the process. In calculating the value of an ore, sulphur -is commonly debited at the rate of 25c. per unit, which represents -approximately the cost of handling and reworking the matte resulting -from it. The practically complete elimination of matte-fall rendered -possible by the H.-H. process may not be, however, an unmixed blessing. -There may be, for example, a small formation of lead sulphide which -causes trouble in the crucible and lead-well, and results in furnace -difficulties and the presentation of a vexatious between-product. - -It may now be attempted to summarize the cost of the converting -process. Assuming the case of an ore assaying lead, 50 per cent.; iron, -15; sulphur, 22; silica, 8, and alumina, etc., 5, let it be supposed -that it is to be fluxed with pure limestone and pure quartz, with the -aim to make a slag containing silica, 30; ferrous oxide, 40; and lime, -20 per cent. A ton of ore will make, in round numbers, 1000 lb. of -slag, and will require 344 lb. of limestone and 130 lb. of quartz, -or we may say roughly one ton of flux must be added to four tons of -ore, wherefore the ore will constitute 80 per cent. of the charge. In -reducing the charge to 3 per cent. sulphur it will lose ultimately -through expulsion of sulphur and carbon dioxide (of the limestone) -about 20 per cent. in weight, wherefore the quantity of material to -be smelted in the blast furnace will be practically equivalent to -the raw sulphide ore in the charge for the roasting furnaces; but in -the roasting furnace the charge is likely to gain weight, because of -the formation of sulphates. Taking the charge, which I have assumed -above, and reckoning that as it comes from the roasting furnace it -will contain 10 per cent. sulphur, all in the form of sulphate, either -of lead or of lime, and that the iron be entirely converted to ferric -oxide, in spite of the expulsion of the carbon dioxide of the limestone -and the combustion of a portion of the sulphur of the ore as sulphur -dioxide, the charge will gain in weight in the ratio of 1:1.19. This, -however, is too high, inasmuch as a portion of the sulphur will remain -as sulphide while a portion of the iron may be as ferrous oxide. The -actual gain in weight will consequently be probably not more than -one-tenth. The following theoretical calculation will illustrate the -changes: - - ─────────────────────┬──────────────────────┬───────────────────────── - RAW CHARGE │ SEMI-ROASTED CHARGE │ FINISHED CHARGE - ─────────────────────┼──────────────────────┼───────────────────────── - {1000 lb. Pb │ {1154 lb. PbO │ { 1154 lb. PbO - { 300 lb. Fe │ { 428 lb. Fe₂O₃ │ { 428 lb. Fe₂O₃(?) - Ore { 160 lb. SiO₂ │ Ore { 160 lb. SiO₂ │ Ore { 160 lb. SiO₂ - { 100 lb. Al₂O₃,│ { 100 lb. Al₂O₃, │ { 100 lb. Al₂O₃, - etc. │ etc. │ etc. - { 440 lb. S │ { 300 lb. S │ { 68 lb. S - │ │ - { 130 lb. SiO₂ │ { 130 lb. SiO₂ │ { 130 lb. SiO₂ - Flux { 344 lb. CaCO₃ │ Flux { 193 lb. CaO │ Flux { 193 lb. CaO - │ 450 lb. O │ - ———— │ ———— │ ———— - 2474 lb. │ 2915 lb. │ 2233 lb. - │ │ - │ 10% S. │ 3% S. - ─────────────────────┴──────────────────────┴───────────────────────── - - Ratios: - - 2474:2915 :: 1:1.18. - 2915:2233 :: 1:0.76⅔. - 2474:2233 :: 1:0.90. - -It may be assumed that for every ton of charge (containing about 80 per -cent. of ore) there will be 1.1 ton of material to go to the converter, -and that the product of the latter will be 0.9 of the weight of the -original charge of raw material. - -Each converter requires 400 cu. ft. of air per minute. The blast -pressure is variable, as different pots are always at different stages -of the process, but assuming the maximum of 16 oz. pressure, with a -blast main of sufficient diameter (at least 15 in.) and the blower -reasonably near the battery of pots, the total requirement is 21 h.p. -The cost of converting will be approximately as follows: - - Labor, 3 foremen at $3.20 $ 9.60 - “ 9 men at $2.50 22.50 - Power, 21 h.p. at 30c 6.30 - Supplies, repairs and renewals 5.00 - —————— - Total $43.40 = 60c. per ton of charge. - -The cost of converting is, of course, reduced directly as the time is -reduced. The above estimate is based on unfavorable conditions as to -time required for working a charge. - -The total cost of treatment, from the initial stage to the delivery of -the desulphurized ore to the blast furnaces, will be, per 2000 lb. of -charge, approximately as follows: - - Crushing 1.0 ton at 10c $0.10 - Mixing 1.0 ton at 10c .10 - Roasting 1.0 ton at 63c .63 - Delivering 1.1 ton to converters at 12c .13 - Converting 1.1 ton at 60c .66 - Breaking 0.9 ton at 60c .54 - ——-——- - Total $2.16 - -The cost per ton of ore will be 2.16 ÷ 0.80 = $2.70. Making allowance -for the crushing of the ore, which is not ordinarily included in the -cost of roasting, and possibly some overestimates, it appears that the -cost of desulphurization by this method, under the conditions assumed -in this paper, is rather higher than in good practice with ordinary -hand-worked furnaces, but it is evident that the cost can be reduced to -approximately the same figure by introduction of improvements, as for -example in breaking the desulphurized ore, and by shortening the time -of converting, which is possible in the case of favorable ores. The -chief advantage must be, however, in the further stage of the smelting. -As to this, there is the evidence that the Broken Hill Proprietary -Company was able to smelt the same quantity of ore in seven furnaces, -after the introduction of the Huntington-Heberlein process, that -formerly required thirteen. A similar experience is reported at -Friedrichshütte, Silesia. - -This increase in the capacity of the blast furnace is due to three -things: (1) In delivering to the furnace a charge containing a reduced -percentage of fine ore, the speed of the furnace is increased, i.e., -more tons of ore can be smelted per square foot of hearth area. (2) -There is less roasted matte to go into the charge. (3) Under some -conditions the percentage of lead in the charge can be increased, -reducing the quantity of gangue that must be fluxed. - -It is difficult to generalize the economy that is effected in the -blast-furnace process, since this must necessarily vary within wide -limits because of the difference in conditions. An increase of 60 to -100 per cent. in blast-furnace capacity does not imply a corresponding -reduction in the cost of smelting. The fuel consumption per ton of ore -remains the same. There is a saving in the power requirements, because -the smelting can be done with a lower blast pressure; also, a saving -in the cost of reworking matte. There will, moreover, be a saving in -other labor, in so far as portions thereof are not already performed -at the minimum cost per ton. The net result under American conditions -of silver-lead smelting can only be determined closely by extensive -operations. That there will be an important saving, however, there is -no doubt. - -The cost of smelting a ton of charge at Denver and Pueblo, exclusive -of roasting and general expense, is about $2.50, of which about $0.84 -is for coke and $1.66 for labor, power and supplies. General expense -amounts to about $0.16 additional. If it should prove possible to -smelt in a given plant 50 per cent. more ore than at present without -increase in the total expense, except for coke, the saving per ton of -charge would be 70c. That is not to be expected, but the half of it -would be a satisfactory improvement. With respect to sulphur in the -charge, the cost is commonly reckoned at 25c. per unit. As compared -with a charge containing 2 per cent. of sulphur there would be a saving -rising toward 50c. per ton as the maximum. It is reasonable to reckon, -therefore, a possible saving of 75c. per ton of charge in silver-lead -smelting, no saving in the cost of roasting, and an increase of about -3 per cent. in the extraction of lead, and perhaps 1 per cent. in the -extraction of silver, as the net results of the application of the -Huntington-Heberlein process in American silver-lead smelting. - -On a charge averaging 12 per cent. lead and 33 oz. silver per ton, -an increase of 3 per cent. in the extraction of lead and 1 per -cent. in the extraction of silver would correspond to 25c. and 35c. -respectively, reckoning lead at 3.5c. per lb., and silver at 60c. per -oz. In this, however, it is assumed that all lead-bearing ores will -be desulphurized by this process, which practically will hardly be -the case. A good deal of pyrites, containing only a little lead, will -doubtless continue to be roasted in Brückner cylinders, and other -mechanical furnaces, which are better adapted to the purpose than are -the lime-roasting pots. Moreover, a certain proportion of high-grade -lead ore, which is now smelted raw, will be desulphurized outside of -the furnace, at additional expense. It is comparatively simple to -estimate the probable benefit of the Huntington-Heberlein process in -the case of smelting works which treat principally a single class of -ore, but in such works as those in Colorado and Utah, which treat a -wide variety of ores, we must anticipate a combination process, and -await results of experience to determine just how it will work out. -It should be remarked, moreover, that my estimates do not take into -account the royalty on the process, which is an actual debit, whether -it be paid on a tonnage basis or be computed in the form of a lump sum -for the license to its use. - -However, in view of the immense tonnage of ore smelted annually for -the extraction of silver and lead, it is evident that the invention of -lime-roasting by Huntington and Heberlein was an improvement of the -first order in the metallurgy of lead. - -In the case of non-argentiferous galena, containing 65 per cent. of -lead (as in southeastern Missouri), comparison may be made with the -slag-roasting and blast-furnace smelting of the ore. Here, no saving -in cost of roasting may be reckoned and no gain in the speed of the -blast furnaces is to be anticipated. The only savings will be in -the increase in the extraction of lead from 92 to 98 per cent., and -the elimination of matte-roasting, which latter may be reckoned as -amounting to 50c. per ton of ore. The extent of the advantage over -the older method is so clearly apparent that it need not be computed -any further. In comparison with the Scotch-hearth bag-house method of -smelting, however, the advantage, if any, is not so certain. That -method already saves 98 per cent. of the lead, and on the whole is -probably as cheap in operation as the Huntington-Heberlein could be -under the same conditions. The Huntington-Heberlein method has replaced -the old roast-reaction method at Tarnowitz, Silesia, but the American -Scotch-hearth method as practised near St. Louis is likely to survive. - -A more serious competitor will be, however, the Savelsberg process, -which appears to do all that the Huntington-Heberlein process does, -without the preliminary roasting. Indeed, if the latter be omitted -(together with its estimated expense of 63c. per ton of charge, or -79c. per ton of ore), all that has been said in this paper as to the -Huntington-Heberlein process may be construed as applying to the -Savelsberg. The charge is prepared in the same way, the method of -operating the converters is the same, and the results of the reactions -in the converters are the same. The litigation which is pending between -the two interests, Messrs. Huntington and Heberlein claiming that -Savelsberg infringes their patents, will be, however, a deterrent to -the extension of the Savelsberg process until that matter be settled. - -The Carmichael-Bradford process may be dismissed with a few words. It -is similar to the Savelsberg, except that gypsum is used instead of -limestone. It is somewhat more expensive because the gypsum has to be -ground and calcined. The process works efficiently at Broken Hill, -but it can hardly be of general application, because gypsum is likely -to be too expensive, except in a few favored localities. The ability -to utilize the converter gases for the manufacture of sulphuric acid -will cut no great figure, save in exceptional cases, as at Broken -Hill, and anyway the gases of the other processes can be utilized for -the same purpose, which is in fact being done in connection with the -Huntington-Heberlein process in Silesia. - -The cost of desulphurizing a ton of galena concentrate by the -Carmichael-Bradford process is estimated by the company controlling -the patents as follows, labor being reckoned at $1.80 per eight hours, -gypsum at $2.40 per 2240 lb., and coal at $8.40 per 2240 lb.: - - 0.25 ton of gypsum $0.60 - Dehydrating and granulating gypsum .48 - Drying mixture of ore and gypsum .12 - Converting 0.24 - Spalling sintered material .12 - 0.01 ton coal .08 - ——-——- - Total $1.64 - -The value of the lime in the sintered product is credited at 12c., -making the net cost $1.52 per 2240 lb. of ore. - -The cost allowed for converting may be explained by the more rapid -action that appears to be attained with the ores of Broken Hill than -with some ores that are treated in North America, but the low figure -estimated for spalling the sintered material appears to be highly -doubtful. - -The theory of the lime-roasting processes is not yet well established. -It is recognized that the explanation offered by Huntington and -Heberlein in their original patent specification is erroneous. There is -no good evidence in their process, or any other, of the formation of -the higher oxide of lime, which they suggest. - -At the present time there are two views. In one, formulated most -explicitly by Professor Borchers, there is formed in this process a -plumbate of calcium, which is an active oxidizing agent. A formation of -this substance was also described by Carmichael in his original patent, -but he considered it to be the final product, not the active oxidizing -agent. - -In the other view, the lime, or limestone, serves merely as a diluent -of the charge, enabling the air to obtain access to the particles of -galena, without liquefaction of the latter. The oxidation of the lead -sulphide is therefore effected chiefly by the air, and the process -is analogous to what takes place in the bessemer converter or in the -Germot process of smelting, or perhaps more closely to what might -happen in an ordinary roasting furnace, provided with a porous hearth, -through which the air supply would be introduced. Roasting furnaces of -that design have been proposed, and in fact such a construction is now -being tested for blende roasting in Kansas. - -Up to the present time, the evidence is surely too incomplete to enable -a definite conclusion to be reached. Some facts may, however, be stated. - -There is clearly reaction to a certain extent between lead sulphide -and lead sulphate, as in the reverberatory smelting furnace, because -prills of metallic lead are to be observed in the lime-roasted charge. - -There is a formation of sulphuric acid in the lime-roasting, upon the -oxidizing effect of which Savelsberg lays considerable stress, since -its action is to be observed on the iron work in which it condenses. - -Calcium sulphate, which is present in all of the processes, being -specifically added in the Carmichael-Bradford, evidently plays an -important chemical part, because not only is the sulphur trioxide -expelled from the artificial gypsum, but also it is to a certain -extent expelled from the natural gypsum, which is added in the -Carmichael-Bradford process; in other words, more sulphur is given off -by the charge than is contained by the metallic sulphides alone. - -Further evidence that lime does indeed play a chemical part in the -reaction is presented by the phenomena of lime-roasting in clay dishes -in the assay muffle, wherein the air is certainly not blown through the -charge, which is simply exposed to superficial oxidation as in ordinary -roasting. - -The desulphurized charge dropped from the pot is certainly at much -below the temperature of fusion, even in the interior, but we have no -evidence of the precise temperature condition during the process itself. - -Pyrite and even zinc blende in the ore are completely oxidized. This, -at least, indicates intense atmospheric action. - -The papers by Borchers,[39] Doeltz,[40] Guillemain,[41] and -Hutchings[42] may profitably be studied in connection with the -reactions involved in lime-roasting. The conclusion will be, however, -that their precise nature has not yet been determined. In view of the -great interest that has been awakened by this new departure in the -metallurgy of lead, it is to be expected that much experimental work -will be devoted to it, which will throw light upon its principles, and -possibly develop it from a mere process of desulphurization into one -which will yield a final product in a single operation. - - - - - PART VI - - OTHER METHODS OF SMELTING - - - - - THE BORMETTES METHOD OF LEAD AND COPPER SMELTING[43] - - BY ALFREDO LOTTI - - (September 30, 1905) - - -It is well known that, in order to obtain a proper fusion in lead -and copper ore-smelting, it is not only advantageous, but often -indispensable, that a suitable proportion of slag be added to the -charge. In the treatment of copper matte in the converter, the total -quantity of slag must be resmelted, inasmuch as it always retains a -notable quantity of the metal; while in the smelting of lead ore in the -blast furnace, the addition of slag is mainly intended to facilitate -the operation, avoiding the use of strong air pressure and thus -diminishing the loss of lead. The proportion of slag required sometimes -amounts to 30 to 35 per cent. of the weight of the ore. - -Inasmuch as the slag is usually added in lump form, cold, its original -heat (about 400 calories per kilogram) is completely lost and an -intimate mixture with the charge cannot be obtained. For this reason, I -have studied the agglomeration of lead and copper ores with fused slag, -employing a variable proportion according to the nature of the ore -treated. In the majority of cases, and with some slight modifications -in each particular case, by incorporating the dry or slightly moistened -mineral with the predetermined quantity of liquid slag, and by rapidly -stirring the mixture so as to secure a proper subdivision of the slag -and the mineral, there is produced a spongy material, largely composed -of small pieces, together with a simultaneous evolution of dense fumes -of sulphur, sulphur dioxide, and sulphur trioxide. By submitting this -spongy material to an air blast, the sulphur of the mineral is burned, -the temperature rising in the interior of the mass to a clear red -heat. Copious fumes of sulphur dioxide and trioxide are given off, -and at times a yellowish vapor of sulphur, which condenses in drops, -especially if the ore is pyritous. - -At the end of from one to three hours, according to the quantity of -sulphur contained in the material under treatment and the amount of the -air pressure, the desulphurization of the ore, so far as it has come -in contact with the air, is completed, and the mass, now thoroughly -agglomerated, forms a spongy but compact block. It is then only -necessary to break it up and smelt it with the requisite quantity of -flux and coke. The physical condition of the material is conducive to -a rapid and economical smelting, while the mixture of the sulphide, -sulphate and oxide leads to a favorable reaction in the furnace. - -In employing this method, it sometimes happens that ores rich in -sulphur produce during the smelting a little more matte than when the -ordinary system of roasting is employed. In such instances, in order -to avoid or to diminish the cost of re-treatment of the matte, it is -best to agglomerate a portion thereof with the crude mineral and the -slag. This has the advantage of oxidizing the matte, which acts as a -ferruginous flux in the smelting. - -The system described above leads to considerable economy, especially -in roasting, as the heat of the scoria, together with that given off -in the combustion of the sulphur, is almost always sufficient for the -agglomeration and desulphurization of the mineral; while, moreover, it -reduces the cost of smelting in the blast furnace. Although the primary -desulphurization is only partial (about 50 per cent.), it continues -in the blast furnace, since the mineral, agglomerated with the slag, -assumes a spongy form and thereby presents an increased surface to the -action of the air. The sulphur also acts as a fuel and does not produce -an excessive quantity of matte. - -The system will prove especially useful in the treatment of -argentiferous lead ore, since, by avoiding the calcination in a -reverberatory furnace, loss of silver is diminished. It appears, -however, that, contrary to the reactions which occur in the -Huntington-Heberlein process, a calcareous or basic gangue is not -favorable to this process, if the proportion be too great. - -The following comparison has been made in the case of an ore containing -62 to 65 per cent. of lead, 16 to 17 per cent. sulphur, 10 to 11 per -cent. zinc, 0.4 per cent. copper, and 0.222 per cent. silver, in which -connection it is to be remarked that, in general, the less zinc there -is in the ore the better are the results. - -[Illustration: FIG. 21.—Elevation and Plan of Converting Chambers.] - -_Ordinary Method._—Roast-reduction. Cost per 1000 kg. of crude ore: - - 1. Roasting in reverberatory furnace: - Labor $0.70 - Fuel 1.50 - Repairs and supplies .05 - ————- $2.25 - - 2. Smelting in water-jacket: - Labor $1.01 - Fuel 2.20 - Repairs and supplies .03 - Fluxes .50 - ————- 3.74 - ————- - Total $5.99 - -_Bormettes Method._—Agglomeration with slag, pneumatic desulphurization -and smelting in water-jacket: - - 1. Agglomeration and desulphurization: - Labor $0.42 - Repairs and supplies 0.05 - ———- $0.47 - - 2. Smelting in water-jacket: - Labor $0.90 - Fuel 1.91 - Repairs and supplies .03 - Fluxes .42 - ————- 3.26 - ————- - Total $3.73 - -This shows a difference in favor of the new method of $2.26 per ton of -ore, without taking into account the savings realized by a much more -speedy handling of the operation, which would further reduce the cost -to approximately $2.50 per ton. - -[Illustration: FIG. 22.—Details of Transfer Cars.] - -In the above figures, no account has been taken of general expenses, -which per ton of ore are reduced because of the greater rapidity of the -process, enabling a larger quantity of ore to be smelted in a given -time. Making allowance for this, the saving will amount to an average -of $2.40 per 1000 kg., a figure which will naturally vary according -to the prices for fuel, labor, and the quantity of matte which it may -be necessary to re-treat. If the quantity of matte does not exceed -10 per cent. of the weight of the ore, it can be desulphurized by -admixture with the ore, without use of other fuel. If, however, the -proportion of matte rises to 20 parts per 100 parts of ore (a maximum -which ought not to be reached in good working), it is necessary to -roast a portion of it. Under unfavorable conditions, consequently, -the saving effected by this process may be reduced to $2 @ $2.20 per -1000 kg., and even to as little as $1.40 @ $1.60. The above reckonings -are, however, without taking any account of the higher extraction of -lead and silver, which is one of the great advantages of the Bormettes -process. - -[Illustration: FIG. 23.—Latest Form of Converter. (Section on A B.)] - -The technical results obtained in the smelting of an ore of the above -mentioned composition are as follows: - - ────────────────────────────────────┬─────────────┬───────────── - │ ORDINARY │ BORMETTES - │ METHOD │ METHOD - ────────────────────────────────────┼─────────────┼───────────── - Coke, per cent. of the charge │ 14 │ 12 - Blast pressure, water gage │12 to 20 cm. │12 to 14 cm. - Tons of charge smelted per 24 hr │ 20 │ 25 - Tons of ore smelted per 24 hr │ 8 │ 10 - Lead assay of slag │0.80 to 0.90%│0.20 to 0.40% - Matte-fall, per cent. of ore charged│ 5 to 10 │ 10 to 15 - Lead extraction │ 90% │ 92% - Silver extraction │ 95% │ 98% - ────────────────────────────────────┴─────────────┴───────────── - -[Illustration: FIG. 24.—Latest Form of Converter. (Section on C D.)] - -The higher extractions of lead and silver are explained by the fact -that the loss of metals in roasting is reduced, while, moreover, the -slags from the blast furnace are poorer than in the ordinary process -of smelting. The economy in coke results from the greater quantity of -sulphur which is utilized as fuel, and from the increased fusibility of -the charge for the blast furnace. - -The new system of desulphurization enables the charge to be smelted -with a less quantity of fresh flux, by the employment in its place of a -greater proportion of foul slag. The reduction in the necessary amount -of flux is due not only to the increased fusibility of the agglomerated -charge, but principally to the fact that in this system the formation -of silicates of lead (which are produced abundantly in ordinary -slag-roasting) is almost nil. It is therefore unnecessary to employ -basic fluxes in order to reduce scorified lead. - -[Illustration: FIG. 25.—Latest Form of Converter. (Plan.)] - -The losses of metal in the desulphurization are less than in the -ordinary method, because the crude mineral remains only a short time -(from one to three hours) in the apparatus for desulphurization and -agglomeration, and the temperature of the process is lower. The -blast-furnace slags are poorer, because there is no formation of -silicate of lead during the agglomeration. - -The Bormettes method, in so far as the treatment of lead ore is -concerned, may be considered a combination process of roast-reaction, -of roast-reduction, and of precipitation-smelting. It is not, however, -restricted to the treatment of lead ore. It may also be applied -to the smelting of pyritous copper-bearing ores. In an experiment -with cupriferous pyrites, containing 20 to 25 per cent. sulphur, it -succeeded in agglomerating and smelting them without use of any fuel -for calcination, effecting a perfect smelting, analogous to pyrite -smelting, with the production of a matte of sufficient degree of -concentration. - -The first cost of plant installation is very much reduced by the -Bormettes method, inasmuch as the ordinary roasting furnaces are almost -entirely dispensed with, apparatus being substituted for them which -cost only one-third or one-fourth as much as ordinary furnaces. The -process presents the advantage, moreover, of being put into immediate -operation, without any expenditure of excess fuel. - -The apparatus required in the process is illustrated in Figs. 21-25. -The apparatus for desulphurization and agglomeration consists of a -cast-iron box, composed of four vertical walls, of which two incline -slightly toward the front. These inclined walls carry the air-boxes. -The other two walls are formed, the one in front by the doors which -give access to the interior, and the other in the rear by a straight -plate. The whole arrangement is surmounted by a hood. The four pieces -when assembled form a box without bottom. Several of these boxes -are combined as a battery. The pots in which the agglomeration and -desulphurization are effected are moved into these boxes on suitable -cars, in the manner shown in the first engraving. A later and more -improved form is shown, however, in Figs. 23-25. - -This process, which is the invention of A. Lotti and has been patented -in all the principal countries, is in successful use at the works of -the Société Anonyme des Mines de Bormettes, at Bormettes, La Londe -(Var), France. Negotiations are now in progress with respect to its -introduction elsewhere in Europe. - - - - - THE GERMOT PROCESS[44] - - BY WALTER RENTON INGALLS - - (November 1, 1902) - - -According to F. Laur, in the _Echo des Mines_ (these notes are -abstracted from _Oest. Zeit._, L., xl, 55, October 4, 1902), A. Germot, -of Clichy, France, made experiments some years ago upon the production -of white lead directly from galena. These led Catelin to attempt the -recovery of metallic lead in a similar way. If air be blown in proper -quantity into a fused mass of lead sulphide the following reaction -takes place: - - 2PbS + 2O = SO₂ + Pb + PbS. - -Thus one-half of the lead is reduced, and it is found collects all the -silver of the ore; the other half is sublimed as lead sulphide, which -is free from silver. The reaction is exothermic to the extent that -the burning of one-half the sulphur of a charge should theoretically -develop sufficient heat to volatilize half of the charge and smelt the -other half. This is almost done in practice with very rich galena, -but not so with poorer ore. The temperature of the furnace must be -maintained at about 1100 deg. C. throughout the whole operation, and -there are the usual losses of heat by radiation, absorption by the -nitrogen of the air, etc. Deficiencies in heat are supplied by burning -some of the ore to white lead, which is mixed with the black fume -(PbS) and by the well-known reactions reduced to metal with evolution -of sulphur dioxide. The final result is therefore the production of -(1) pig lead enriched in silver; (2) pig lead free from silver; (3) a -leady slag; and (4) sulphur dioxide. In the case of ores containing -less than 75 per cent. Pb the gangue forms first a little skin and -then a thick hard crust which soon interferes with the operation, -especially if the ore be zinkiferous. This difficulty is overcome by -increasing the temperature or by fluxing the ore so as to produce a -fusible slag. A leady slag is always easily produced; this is the only -by-product of the process. The theoretical reaction requires 600 cu. m. -of air, assuming a delivery of 50 per cent. from the blower, and at one -atmosphere pressure involves the expenditure of 18 h.p. per 1000 kg. of -galena per hour. - -[Illustration: FIG. 26.—Plan and Elevation of Smelting Plant at Clichy.] - -The arrangement of the plant at Clichy is shown diagrammatically in -Fig. 26. There is a round shaft furnace, 0.54 meter in diameter and -4.5 meters high. Power is supplied to the blower C through the pulley -G and the shaft DD. The compressed air is accumulated in the reservoir -R, whence it is conducted by the pipe to the tuyere which is suspended -inside of the furnace by means of a chain, whereby it can be raised -or lowered. O₁ and O₂ are tap-holes. L is a door and N an -observation tube. A is the charge tube. X is the pipe which conveys the -gas and fume to the condensation chambers. T is the pipe through which -the waste gases are drawn. V is the exhauster and S is the chimney. -K₁ and K₂ are tilting crucible furnaces for melting lead and -galena. - -After the furnace has been properly heated, 100 kg. of lead melted in -K₁ are poured in through the cast-iron pipe P, and after that about -200 kg. of pure, thoroughly melted galena from K₂. Ore containing 70 -to 80 per cent. Pb must be used for this purpose. The blast of air is -then introduced into the molten galena, and from 1000 to 3000 kg. of -ore is gradually charged in through the tube A. During this operation -black fume (PbS) collects in the condensation chamber. All outlets are -closed against the external air. If the air blast is properly adjusted, -nothing but black fume is produced; if it begins to become light -colored, charging is discontinued and the blast of air is shut off. -Lead is then tapped through O₂, which is about 0.2 meter above the -hearth, so there is always a bath of lead in the bottom of the furnace; -but it is advisable now and then to tap off some through O₁, so as -gradually to heat up the bottom of the furnace. Hearth accretions are -also removed through O₁. The lead is tapped off through O₂ until -matte appears. The tap hole is then closed, the tuyere is lowered and -the blast is turned into the lead in order to oxidize it and completely -desulphurize the sulphur combinations, which is quickly done. The -oxide of lead is scorified as a very fusible slag, which is tapped off -through O₂, and more ore is then charged in upon the lead bath and -the cycle of operations is begun again. - - - - - PART VII - - DUST AND FUME RECOVERY - - FLUES, CHAMBERS AND BAG-HOUSES - - - - - DUST CHAMBER DESIGN - - BY MAX J. WELCH - - (September 1, 1904) - - -Only a few years ago smelting companies began to recognize the -advantage of large chambers for collecting flue dust and condensing -fumes. The object is threefold: First, profit; second, to prevent law -suits with surrounding agricultural interests; third, cleanliness about -the plant. It is my object at present to discuss the materials used in -construction and general types of cross-section. - -Most of the old types of chambers are built after one general pattern, -namely, brick or stone side walls and arch roof, with iron buckstays -and tie rods. The above type is now nearly out of use, because it is -short-lived, expensive, and dangerous to repair, while the steel and -masonry are not used to good advantage in strength of cross-section. - -With the introduction of concrete and expanded metal began a new -era of dust-chamber construction. It was found that a skeleton of -steel with cement plaster is very strong, light and cheap. The first -flue of the type shown in Fig. 29 was built after the design of E. -H. Messiter, at the Arkansas Valley smelter in Colorado. This flue -was in commission several years, conveying sulphurous gases from the -reverberatory roaster plant. The same company decided, in 1900, to -enlarge and entirely rebuild its dust-chamber system, and three types -of cross-section were adopted to meet the various conditions. All three -types were of cement and steel construction. - -The first type, shown in Fig. 27, is placed directly behind the blast -furnaces. The cross-section is 273 sq. ft. area, being designed for a -10-furnace lead smelter. The back part is formed upon the slope of the -hillside and paved with 2.5 in. of brick. The front part is of ribbed -cast-iron plates. Ninety per cent. of the flue dust is collected in -this chamber and is removed, through sliding doors, into tram cars. -There is a little knack in designing a door to retain flue dust. It is -simply to make the bottom sill of the door frame horizontal for a space -of about 1 in. outside of the door slide. - -The front part of the chamber, Fig. 27, is of expanded metal and -cement. The top is of 20 in. I-beams, spanning a distance of 24 ft. -with 15 in. cross-beams and 3 in. of concrete floor resting upon -the bottom flanges of the beams. This heavy construction forms the -foundation for the charging floor, bins, scales, etc. - -[Illustration: FIG. 27.—Rectangular form of Concrete Dust Chamber.] - -While dwelling upon this type of construction I wish to mention a -most important point, that of the proper factor of safety. Flue dust, -collected near the blast furnace, weighs from 80 to 100 lb. per cubic -foot, and the steel supports should be designed for 16,000 lb. extreme -fiber stress, when the chamber is three-quarters full of dust. If the -dust is allowed to accumulate beyond this point, the steel, being well -designed, should not be overstrained. Discussions as to strains in -bins have been aired by the engineering profession, but the present -question is “Where is a dust chamber a bin?” Experience shows that bin -construction should be adopted behind, or in close proximity to, the -blast furnaces. - -Fig. 28 shows the second type of hopper-bottom flue adopted. It is -of very light construction, of 274 sq. ft. area in the clear. The -beginning of this flue being 473 ft. from the blast furnaces removes -all possibility of any material floor-load, as the dust is light in -weight and does not collect in large quantities. The hopper-bottom -floor is formed of 4 in. concrete slabs, in panels, placed between 4 -in. I-beams. Cast-iron door frames, with openings 12 × 16 in., are -placed on 5 ft. centers. The concrete floor is tamped in place around -the frames. The side walls and roof are built of 1 in. angles, expanded -metal, and plastered to 2.5 in. thickness. At every 10 ft. distance, -pilaster ribs built of 2 in. angles, latticed and plastered, form the -wind-bracing and arch roof support. - -[Illustration: FIG. 28.—Arched form of Concrete Dust Chamber.] - -Fig. 29 shows the beehive construction. This chamber is of 253 sq. ft. -cross-sectional area. It is built of 2 in. channels, placed 16 in. -centers, tied with 1 × 0.125 in. steel strips. The object of the strips -is to support the 2 in. channels during erection. No. 27 gage expanded -metal lath was wired to the inside of the channels and the whole -plastered to a thickness of 3 in. The inside coat was plastered first -with portland cement and sand, one to three, with about 5 per cent. -lime. The filling between ribs is one to four, and the outside coat one -to three. - -The above types of dust chamber have been in use over three years at -Leadville. Cement and concrete, in conjunction with steel, have been -used in Utah, Montana and Arizona, in various types of cross-section. -The results show clearly where not to use cement; namely, where -condensing sulphur fumes come in contact with the walls, or where -moisture collects, forming sulphuric acid. The reason is that portland -cement and lime mortar contain calcium hydrate, which takes up sulphur -from the fumes, forming calcium sulphate. In condensing chambers, this -calcium sulphate takes up water, forming gypsum, which expands and -peels off. - -[Illustration: FIG. 29.—Beehive form of Concrete Dust Chamber.] - -In materials of construction it is rather difficult to get something -that will stand the action of sulphur fumes perfectly. The lime mortar -joints in the old types of brick flues are soon eaten away. The arches -become weak and fall down. I noted a sheet steel condensing system, -where in one year the No. 12 steel was nearly eaten through. With a -view of profiting by past experience, let us consider the acid-proof -materials of construction, namely, brick, adobe mortar, fire-clay, and -acid-proof paint. Also, let us consider at what place in a dust-chamber -system are we to take the proper precaution in the use of these -materials. - -At smelting plants, both copper and lead, it is found that near the -blast furnaces the gases remain hot and dry, so that concrete, brick -or stone, or steel, can safely be used. Lead-blast furnace gases will -not injure such construction at a distance of 6 or 8 ft. away from the -furnaces. For copper furnaces, roasters or pyritic smelting, concrete -or lime mortar construction should be limited to within 200 or 300 ft. -of the furnaces. - -Another type of settling chamber is 20 ft. square in the clear, with -concrete floor between beams and steel hopper bottom. This chamber -is built within 150 ft. distance from the blast furnaces, and is one -of the types used at the Shannon Copper Company’s plant at Clifton, -Arizona. After passing the 200 ft. mark, there is no need of expensive -hopper design. The amount of flue dust settled beyond this point is so -small that it is a better investment to provide only small side doors -through which the dust can be removed. The ideal arrangement is to have -a system of condensing chambers, so separated by dampers that either -set can be thrown out for a short time for cleaning purposes, and the -whole system can be thrown in for best efficiency. - -As to cross-section for condensing chambers, I consider that the -following will come near to meeting the requirements. One, four, and -six, concrete foundation; tile drainage; 9 in. brick walls, laid in -adobe mortar, pointed on the outside with lime mortar; occasional -strips of expanded metal flooring laid in joints; the necessary -pilasters to take care of the size of cross-section adopted; the top -covered with unpainted corrugated iron, over which is tamped a concrete -roof, nearly flat; concrete to contain corrugated bars in accordance -with light floor construction; and lastly, the corrugated iron to have -two coats of graphite paint on under side. - -The above type of roof is used under slightly different conditions over -the immense dust chamber of the new Copper Queen smelter at Douglas, -Arizona. The paint is an important consideration. Steel work imbedded -in concrete should never be painted, but all steel exposed to fumes -should be covered by graphite paint. Tests made by the United States -Graphite Company show that for stack work the paint, when exposed to -acid gases, under as high a temperature as 700 deg. F., will wear well. - - - - - CONCRETE IN METALLURGICAL CONSTRUCTION[45] - - BY HENRY W. EDWARDS - - -The construction of concrete flues of the section shown in Fig. 31 -gives better results than that shown in Fig. 30, being less liable -to collapse. It costs somewhat more to build owing to the greater -complication of the crib, which, in both cases, consists of an interior -core only. For work 4 in. in thickness and under, I recommend the -use of rock or slag crushed to pass through a 1.5 in. ring. Although -concrete is not very refractory, it will easily withstand the heat -of the gases from a set of ordinary lead-or copper-smelting blast -furnaces, or from a battery of calcining or roasting furnaces. I have -never noticed that it is attacked in any way by sulphur dioxide or -other furnace gas. - -[Illustration: FIGS. 30 and 31.—Sections of Concrete Flues.] - -Shapes the most complicated to suit all tastes in dust chambers can -be constructed of concrete. The least suitable design, so far as the -construction itself is concerned, is a long, wide, straight-walled, -empty chamber, which is apt to collapse, either inwards or outwards, -and, although the outward movement can be prevented by a system of -light buckstays and tie-rods, the tendency to collapse inwards is -not so simply controlled in the absence of transverse baffle walls. -The tendency, so far as the collection of mechanical flue dust is -concerned, appears to be towards a large empty chamber, without -baffles, etc., in which the velocity of the air currents is reduced to -a minimum, and the dust allowed to settle. In the absence of transverse -baffle walls to counteract the collapsing tendency, it seems best to -design the chamber with a number of stout concrete columns at suitable -intervals along the side and end walls—the walls themselves being made -only a few inches thick with woven-wire screen or “expanded metal” -buried within them. The wire skeleton should also be embedded into the -columns in order to prevent the separation of wall and the columns. -This method of constructing is one that I have followed with very -satisfactory results as far as the construction itself is concerned. - -[Illustration: FIG. 32.—Concrete Dust Chamber at the Guillermo Smelting -Works, Palomares, Spain. (Horizontal section.)] - -Figs. 32 and 33 show a chamber designed and erected at the Don -Guillermo Smelting Works at Palomares, Province of Murcia, Spain. -Figs. 34 and 35 show a design for the smelter at Murray Mine, Sudbury, -Ontario, in which the columns are hollow, thus economizing concrete -material. For work of this kind the columns are built first and the -wire netting stretched from column to column and partly buried within -them. The crib is then built on each side of the netting, a gang of men -working from both sides, and is built up a yard or so at a time as the -work progresses. Doors of good size should be provided for entrance -into the chamber, and as they will seldom be opened there is no need -for expensive fastenings or hinges. - -[Illustration: FIG. 33.—Concrete Dust Chamber at the Guillermo Smelting -Works, Palomares, Spain. (End elevation.)] - -_Foundations for Dynamos and other Electrical Machinery._—Dry concrete -is a poor conductor of electricity, but when wet it becomes a fairly -good conductor. Therefore, if it be necessary to insulate the -electrical apparatus, the concrete should be covered with a layer of -asphalt. - -[Illustration: FIG. 34.—Concrete Dust Chamber designed for smelter at -Murray Mine, Sudbury, Ontario, Can. There are eight 9 ft. sections in -the plan.] - -_Chimney Bases._—Fig. 36 shows the base for the 90 ft. brick-stack at -Don Guillermo. The resemblance to masonry is given by nailing strips of -wood on the inside of the crib. - -[Illustration: FIG. 35.—Concrete Dust Chamber designed for smelter at -Murray Mine, Sudbury, Ontario, Can. (End elevation.)] - -_Retaining-Walls._—Figs. 37, 38, and 39 show three different styles -of retaining-walls, according to location. These walls are shown -in section only, and show the placing of the iron reenforcements. -Retaining-walls are best built in panels (each panel being a day’s -work), for the reason that horizontal joints in the concrete are -thereby avoided. The alternate panels should be built first and the -intermediate spaces filled in afterward. Should there be water behind -the wall it is best to insert a few small pipes through the wall, in -order to carry it off; this precaution is particularly important in -places where the natural surface of the ground meets the wall, as -shown in Figs. 37 and 38. If a wooden building is to be erected on the -retaining-wall, it is best to bury a few 0.75 in. bolts vertically in -the top of the wall, by which a wooden coping may be secured (see Figs. -37, 38, and 39), which forms a good commencement for the carpenter work. - -[Illustration: FIG. 36.—Concrete Base for a 90 ft. Chimney at the -Guillermo Smelting Works, Palomares, Spain.] - -Minimum thickness for a retaining-wall, having a liberal quantity -of iron embedded therein, is 20 in. at the bottom and 10 in. at the -top, with the taper preferably on the inner face. In the absence of -interior strengthening irons the thickness of the wall at the bottom -should never be less than one-fourth the total hight, and at the top -one-seventh of the hight; unless very liberal iron bracing be used, -the dimensions can hardly be reduced to less than one-seventh and -one-tenth respectively. Unbraced retaining-walls are more stable with -the batter on the outer face. Dry clay is the most treacherous material -that can be had behind a retaining-wall, especially if it be beaten -in, for the reason that it is so prone to absorb moisture and swell, -causing an enormous side thrust against the wall. When this material is -to be retained it is best to build the wall superabundantly strong—a -precaution which applies even to a dry climate, because the bursting -of a water-pipe may cause the damage. In order to avoid horizontal -joints it is best, wherever practicable, to build the crib-work in its -entirety before starting the concrete. In a retaining-wall 3 ft. thick -by 16 ft. high this is not practicable. The supporting posts and struts -can, however, be completed and the boards laid in as the wall grows, -in order not to interrupt the regular progress of the tamping. A good -finish may be produced on the exposed face of the wall by a few strokes -of the shovel up and down with its back against the crib. - -[Illustration: FIGS. 37, 38, and 39.—Retaining-Walls of Concrete.] - -In conclusion I wish to state that this paper is not written for the -instruction of the civil engineer, or for those who have special -experience in this line; but rather for the mining engineer or -metallurgist whose training is not very deep in this direction, and who -is so often thrown upon his own resources in the wilderness, and who -might be glad of a few practical suggestions from one who has been in a -like predicament. - - - - - CONCRETE FLUES[46] - - BY EDWIN H. MESSITER - - (September, 1904) - - -Under the heading “Flues,” Mr. Edwards refers to the Beehive -construction, a cross-section of which is shown in Fig. 31 of his -paper. A flue similar to this was designed by me about six years -ago,[47] and in which the walls, though much thinner than those -described by Mr. Edwards, gave entire satisfaction. These walls, from -2.25 in. thick throughout in the smaller flues to 3.25 in. in the -larger, were built by plastering the cement mortar on expanded-metal -lath, without the use of any forms or cribs whatever, at a cost of -labor generally less than $1 per sq. yd. of wall. Of course, where -plasterers cannot be obtained on reasonable terms, the cement can be -molded between wooden forms, though it is difficult to see how it can -be done with an interior core only, as stated by Mr. Edwards. - -In regard to the effect of sulphur dioxide and furnace gases on the -cement, I have found that in certain cases this is a matter which -must be given very careful attention. Where there is sufficient heat -to prevent the existence of condensed moisture inside of the flue, -there is apparently no action whatever on the cement, but if the -concrete is wet, it is rapidly rotted by these gases. At points near -the furnaces there is generally sufficient heat not only to prevent -internal condensation of the aqueous vapor always present in the gases, -but also to evaporate water from rain or snow falling on the outside -of the flue. Further along a point is reached where rain-water will -percolate through minute cracks caused by expansion and contraction, -and reach the interior even though internal condensation does not occur -there in dry weather. From this point to the end of the flue the roof -must be coated on the outside with asphalt paint or other impervious -material. In very long flues a point may be reached where moisture will -condense on the inside of the walls in cold weather. From this point -to the end of the flue it is essential to protect the interior with an -acid-resisting paint, of which two or more coats will be necessary. -For the first coat a material containing little or no linseed oil is -best, as I am informed that the lime in the cement attacks the oil. For -this purpose I have used ebonite varnish, and for the succeeding coats -durable metal-coating. The first coat will require about 1 gal. of -material for each 100 sq. ft. of surface. - -In one of the earliest long flues built of cement in this country, a -small part near the chimney was damaged as a result of failure to apply -the protective coating, the necessity for it not having been recognized -at the time of its construction. It may be said, in passing, that other -long brick flues built prior to that time were just as badly attacked -at points remote from the furnaces. In order to reduce the amount of -flue subject to condensation, the plastered flues have been built with -double lath having an intervening air-space in the middle of the wall. - -In building thin walls of cement, such as flue walls, it is -particularly important to prevent them from drying before the cement -has combined with all the water it needs. For this reason the work -should be sprinkled freely until the cement is fully set. Much work of -this class has been ruined through ignorance by fires built near the -walls in cold weather, which caused the mortar to shell off in a short -time. - -The great saving in cost of construction, which the concrete-steel flue -makes possible, will doubtless cause it to supersede other types to -even a greater extent than it has already done. If properly designed -this type of construction reduces the cost of flues by about one-half. -Moreover, the concrete-steel flue is a tight flue as compared with -one built of brick. There is a serious leakage through the walls of -the brick flues which is not easily observed in flues under suction -as most flues are, but when a brick flue is under pressure from a -fan the leakage is surprisingly apparent. In flues operating by -chimney-draft the entrance of cold air must cause a considerable loss -in the efficiency of the chimney, a disadvantage which would largely be -obviated by the use of the concrete-steel flue. - - - - - CONCRETE FLUES[48] - - BY FRANCIS T. HAVARD - - -In discussion of Mr. Edwards’s interesting and valuable paper, I -beg to submit the following notes concerning the advantages and -disadvantages of the concrete flues and stacks at the plant of the -Anhaltische Blei-und Silber-werke. The flues and smaller stacks at the -works were constructed of concrete consisting generally of one part of -cement to seven parts of sand and jig-tailings but, in the case of the -under-mentioned metal concrete slabs, of one part of cement to four -parts of sand and tailings. The cost of constructing the concrete flue -approximated 5 marks per sq. m. of area (equivalent to $0.11 per sq. -ft.). - -_Effect of Heat._—A temperature above 100 deg. C. caused the concrete -to crack destructively. Neutral furnace gases at 120 deg. C., passing -through an independent concrete flue and stack, caused so much damage -by the formation of cracks that, after two years of use, the stack, -constructed of pipes 4 in. thick, required thorough repairing and -auxiliary ties for every foot of hight. - -_Effect of Flue Gases and Moisture._—The sides of the main flue, made -of blocks of 6 in. hollow wall-sections, 100 cm. by 50 cm. in area, -were covered with 2 in. or 1 in. slabs of metal concrete. In cases -where the flue was protected on the outside by a wooden or tiled roof, -and inside by an acid-proof paint, consisting of water-glass and -asbestos, the concrete has not been appreciably affected. In another -case, where the protective cover, both inside and outside, was of -asphalt only, the concrete was badly corroded and cracked at the end -of three years. In a third case, in which the concrete was unprotected -from both atmospheric influence on the outside, and furnace gases on -the inside, the flue was quite destroyed at the end of three years. -That portion of the protected concrete flue, near the main stack, which -came in contact only with dry, cold gases was not affected at all. - -Gases alone, such as sulphur dioxide, sulphur trioxide, and others, -do not affect concrete; neither is the usual quantity of moisture -in furnace gases sufficient to damage concrete; but should moisture -penetrate from the outside of the flue, and, meeting gaseous SO₂ or -SO₃, form hydrous acids, then the concrete will be corroded. - -_Effect of the Atmosphere Alone._—For outside construction work, -foundations and other structures not exposed to heat, moist acid gases -and chemicals, the concrete has maintained its reputation for cheapness -and durability. - -_Effect of Crystallization of Contained Salts._—In chemical works, -floors constructed of concrete are sometimes unsatisfactory, for the -reason that soluble salts, noticeably zinc sulphate, will penetrate -into the floor and, by crystallizing in narrow confines, cause the -concrete to crack and the floor to rise in places. - - - - - BAG-HOUSES FOR SAVING FUME - - BY WALTER RENTON INGALLS - - (July 15, 1905) - - -One of the most efficient methods of saving fume and very fine dust in -metallurgical practice is by filtration through cloth. This idea is by -no means a new one, having been proposed by Dr. Percy, in his treatise -on lead, page 449, but he makes no mention of any attempt to apply it. -Its first practical application was found in the manufacture of zinc -oxide direct from ores, initially tried by Richard and Samuel T. Jones -in 1850, and in 1851 modified by Samuel Wetherill into the process -which continues in use at the present time in about the same form as -originally. In 1878 a similar process for the manufacture of white -lead direct from galena was introduced at Joplin, Mo., by G. T. Lewis -and Eyre O. Bartlett, the latter of whom had previously been engaged -in the manufacture of zinc oxide in the East, from which he obtained -his idea of the similar manufacture of white lead. The difference -in the character of the ore and other conditions, however, made it -necessary to introduce numerous modifications before the process became -successful. The eventual success of the process led to its application -for filtration of the fume from the blast furnaces at the works of the -Globe Smelting and Refining Company, at Denver, Colo., and later on for -the filtration of the fume from the Scotch hearths employed for the -smelting of galena in the vicinity of St. Louis. - -In connection with the smelting of high-grade galena in Scotch hearths, -the bag-house is now a standard accessory. It has received also -considerable application in connection with silver-lead blast-furnace -smelting and in the desilverizing refineries. Its field of usefulness -is limited only by the character of the gas to be filtered, it being -a prerequisite that the gas contain no constituent that will quickly -destroy the fabric of which the bags are made. Bags are also employed -successfully for the collection of dust in cyanide mills, and other -works in which fine crushing is practised, for example, in the -magnetic separating works of the New Jersey Zinc Company, Franklin, -N. J. , where the outlets of the Edison driers, through which the ore -is passed, communicate with bag-filtering machines, in which the bags -are caused to revolve for the purpose of mechanical discharge. The -filtration of such dust is more troublesome than the filtration of -furnace fume, because the condensation of moisture causes the bags to -become soggy. - -[Illustration: FIG. 40.—Bag-house, Globe Smelting Works.] - -The standard bag-house employed in connection with furnace work is a -large room, in which the bags hang vertically, being suspended from -the top. The bags are simply tubes of cotton or woolen (flannel) -cloth, from 18 to 20 in. in diameter, and 20 to 35 ft. in length, most -commonly about 30 ft. In the manufacture of zinc oxide, the fume-laden -gas is conducted into the house through sheet-iron pipes, with suitably -arranged branches, from nipples on which the bags are suspended, the -lower end of the bag being simply tied up until it is necessary to -discharge the filtered fume by shaking. In the bag-houses employed in -the metallurgy of lead, the fume is introduced at the bottom into brick -chambers, which are covered with sheet-iron plates, provided with the -necessary nipples; or else into hopper-bottom, sheet-iron flues, with -the necessary nipples on top. In either case the bags are tied to the -nipples, and are tied up tight at the top, where they are suspended. -When the fume is dislodged by shaking the bags, it falls into the -chamber or hopper at the bottom, whence it is periodically removed. - -The cost of attending a bag-house, collecting the fume, etc., varies -from about 10c. per ton of ore smelted in a large plant like the Globe, -to about 25c. per ton in a Scotch-hearth plant treating 25 tons of ore -per 24 hours. - -No definite rules for the proportioning of filtering area to the -quantity of ore treated have been formulated. The correct proportion -must necessarily vary according to the volume of gaseous products -developed in the smelting of a ton of ore, the percentage of dust and -fume contained, and the frequency with which the bags are shaken. -It would appear, however, that in blast furnaces and Scotch-hearth -smelting a ratio of 1000 sq. ft. per ton of ore would be sufficient -under ordinary conditions. The bag-house originally constructed at -the Globe works had about 250 sq. ft. of filtering area per ton of -charge smelted, but this was subsequently increased, and Dr. Iles, -in his treatise on lead-smelting, recommends an equipment which would -correspond to about 750 sq. ft. per ton of charge. At the Omaha works, -where the Brown-De Camp system was used, there was 80,000 sq. ft. of -cloth for 10 furnaces 42 × 120 in., according to Hofman’s “Metallurgy -of Lead,” which would give about 1000 sq. ft. per ton of charge -smelted, assuming an average of eight furnaces to be in blast. A -bag-house in a Scotch-hearth smeltery, at St. Louis, had approximately -900 sq. ft. per ton of ore smelted. At the Lone Elm works, at Joplin, -the ratio was about 3500 sq. ft. per ton of ore smelted, when the -works were run at their maximum capacity. In the manufacture of zinc -oxide the bag area used to be from 150 to 200 sq. ft. per square foot -of grate on which the ore is burned, but at Palmerton, Pa. (the most -modern plant), the ratio is only 100:1. This corresponds to about 1400 -sq. ft. of bag area per 2000 lb. of charge worked on the grate. In the -manufacture of zinc-lead white at Cañon City, Colo., the ratio between -bag area and grate area is 150:1. - -Assuming the gas to be free, or nearly free, from sulphurous fumes, the -bags are made of unbleached muslin, varying in weight from 0.4 to 0.7 -oz. avoirdupois per square foot. The cloth should have 42 to 48 threads -per linear inch in the warp and the same number in the woof. A kind of -cloth commonly used in good practice weighs 0.6 oz. per square foot and -has 46 threads per linear inch in both the warp and the woof. - -The bags should be 18 to 20 in. in diameter. Therefore the cloth should -be of such width as to make that diameter with only one seam, allowing -for the lap. Cloth 62 in. in width is most convenient. It costs 4 to -5c. per yard. The seam is made by lapping the edges about 1 in., or -by turning over the edges and then lapping, in the latter case the -stitches passing through four thicknesses of the cloth. It should be -sewed with No. 50 linen thread, making two rows of double lock-stitches. - -The thimbles to which the bags are fastened should be of No. 10 sheet -steel, the rim being formed by turning over a ring of 0.25 in. wire. -The bags are tied on with 2 in. strips of muslin. The nipples are -conveniently spaced 27 in. apart, center to center, on the main pipe. - -The gas is best introduced at a temperature of 250 deg. F. Too high -a temperature is liable to cause them to ignite. They are safe at 300 -deg. F., but the temperature should not be allowed to exceed that point. - -The gas is cooled by passage through iron pipes of suitable radiating -surface, but the temperature should be controlled by a dial thermometer -close to the bag-house, which should be observed at least hourly, and -there should be an inlet into the pipe from the outside, so that, in -event of rise of temperature above 300 deg., sufficient cold air may be -admitted to reduce it within the safety limit. - -In the case of gas containing much sulphur dioxide, and especially any -appreciable quantity of the trioxide, the bags should be of unwashed -wool. Such gas will soon destroy cotton, but wool with the natural -grease of the sheep still in it is not much affected. The gas from -Scotch hearths and lead-blast furnaces can be successfully filtered, -but the gas from roasting furnaces contains too much sulphur trioxide -to be filtered at all, bags of any kind being rapidly destroyed. - - - - - PART VIII - - BLOWERS AND BLOWING ENGINES - - - - - ROTARY BLOWERS VS. BLOWING ENGINES FOR LEAD SMELTING - - (April 27, 1901) - - -A note in the communication from S. E. Bretherton on “Pyritic Smelting -and Hot Blast,” published in the _Engineering and Mining Journal_ -of April 13, 1901, refers to a subject of great interest to lead -smelters. Mr. Bretherton remarked that he had been recently informed -by August Raht that by actual experiment the loss with the ordinary -rotary blowers was 100 per cent. under 10 lb. pressure; that is, it -was possible to shut all the gates so that there was no outlet for the -blast to escape from the blower and the pressure was only 10 lb., or in -other words the blower would deliver no air against 10 lb. pressure. -For that reason Mr. Raht expressed himself as being in favor of blowing -engines for lead blast furnaces. This is of special interest, inasmuch -as it comes from one who is recognized as standing in the first rank of -lead-smelting engineers. Mr. Raht is not alone in holding the opinion -he does. - -The rotary blower did good service in the old days when the air was -blown into the lead blast furnace at comparatively moderate pressure. -At the present time, when the blast pressure employed is commonly -40 oz. at least, and sometimes as high as 48 oz., the deficiencies -of the rotary blower have become more apparent. Notwithstanding the -excellent workmanship which is put into them by their manufacturers, -the extensive surfaces of contact are inherent to the type, and -leakage of air backward is inevitable and important at the pressures -now prevailing. The impellers of a rotary blower should not touch -each other nor the cylinders in which they revolve, but they are made -with as little clearance as possible, the surfaces being coated with -grease, which fills the clearance space and forms a packing. This -will not, however, entirely prevent leakage, which will naturally -increase with the pressure. Even the manufacturers of rotary blowers -admit the defects of the type, and concede that for pressures of 5 -lb. and upward the cylinder blowing engine is the more economical. -Metallurgists are coming generally to the opinion, however, that -blowing engines are probably more economical for pressures of 4 lb. or -thereabouts, and some go even further. With the blowing engines the -air-joints of piston and cylinder are those of actual contact, and -the metallurgist may count on his cubic feet of air, whatever be the -pressure. Blowing engines were actually introduced several years ago -by M. W. Iles at what is now the Globe plant of the American Smelting -and Refining Company, and we believe their performance has been found -satisfactory. - -The fancied drawback to the use of blowing engines is their greater -first cost, but H. A. Vezin, a mechanical engineer whose opinions carry -great weight, pointed out five years ago in the _Transactions_ of the -American Institute of Mining Engineers (Vol. XXVI) that per cubic -foot of air delivered the blowing engine was probably no more costly -than the rotary blower, but on the contrary cheaper, stating that the -first cost of a cylinder blower is only 20 to 25 per cent. more than -that of a rotary blower of the same nominal capacity and the engine -to drive it. The capacity of a rotary blower is commonly given as the -displacement of the impellers per revolution, without allowance for -slip or leakage backward. Mr. Vezin expressed the opinion that for the -same actual capacity at 2 lb. pressure, that is, the delivery in cubic -feet against 2 lb. pressure, the cylinder blower would cost no more -than, if as much as, the rotary blower. - -In this connection it is worth while making a note of the increasing -tendency of lead smelters to provide much more powerful blowers than -were formerly considered necessary, due, no doubt, in large measure to -the recognition of the greater loss of air by leakage backward at the -pressure now worked against. It is considered, for example, that a 42 × -140 in. furnace to be driven under 40 oz. pressure should be provided -with a No. 10 blower, which size displaces 300 cu. ft. of air per -revolution and is designed to be run at about 100 r.p.m.; its nominal -capacity is, therefore, 30,000 cu. ft. of air per minute; although -its actual delivery against 40 oz. pressure is much less, as pointed -out by Mr. Raht and Mr. Bretherton. The Connersville Blower Company, -of Connersville, Ind., lately supplied the Aguas Calientes plant (now -of the American Smelting and Refining Company) with a rotary blower -of the above capacity, and duplicates of it have been installed at -other smelting works. The force required to drive such a huge blower -is enormous, being something like 400 h.p., which makes it advisable -to provide each blower with a directly connected compound condensing -engine. - -In view of the favor with which cylindrical blowing engines for driving -lead blast furnaces are held by many of the leading lead-smelting -engineers, and the likelihood that they will come more and more into -use, it will be interesting to observe whether the lead smelters will -take another step in the tracks of the iron smelters and adopt the -circular form of blast furnace that is employed for the reduction -of iron ore. The limit of size for rectangular furnaces appears to -have been reached in those of 42 × 145 in., or approximately those -dimensions. A furnace of 66 × 160 in., which was built several years -ago at the Globe plant at Denver, proved a failure. H. V. Croll at -that time advocated the building of a circular furnace instead of the -rectangular furnace of those excessive dimensions and considered that -the experience with the latter demonstrated their impracticability. In -the _Engineering and Mining Journal_ of May 28, 1898, he stated that -there was no good reason, however, why a furnace of 300 to 500 tons -daily capacity could not be run successfully, but considered that the -round furnace was the only form permissible. We are unaware whether -Mr. Croll was the first to advocate the use of large circular furnaces -for lead smelting, but at all events there are other experienced -metallurgists who now agree with him, and the time is, perhaps, not far -distant when they may be adopted. - - - - - ROTARY BLOWERS VS. BLOWING ENGINES - - BY J. PARKE CHANNING - - (June 8, 1901) - - -In the issues of the _Engineering and Mining Journal_ for April -13th and 27th reference was made to the relative efficiency of -piston-blowing engines and rotary blowers of the impeller type, and in -these articles August Raht was quoted as saying that, with an ordinary -rotary blower working against 10 lb. pressure, the loss was 100 per -cent. I have waited some time with the idea that some of the blower -people would call attention to the concealed fallacy in the statement -quoted, but so far have failed to notice any reference to the matter. I -feel quite sure that Mr. Bretherton failed to quote Mr. Raht in full. -The one factor missing in this statement is the speed at which the -blower was run when the loss was 100 per cent. - -The accepted method of testing the volumetric efficiency of rotary -blowers is that of “closed discharge.” The discharge opening of the -blower is closed, a pressure gage is connected with the closed delivery -pipe, and the blower is gradually speeded up until the gage registers -the required pressure. The number of revolutions which the blower -makes while holding that pressure, multiplied by the cubic feet per -revolution, will give the total slip of that particular blower at that -particular pressure. Experience has shown that, within the practical -limits of speed at which a blower is run, the slip is a function of -the pressure and has nothing to do with the speed. If, therefore, it -were found that the particular blower referred to by Mr. Raht were -obliged to be revolved at the rate of 30 r.p.m. in order to maintain a -constant pressure of 10 lb. with a closed discharge, and if the blower -were afterward put in practical service, delivering air, and were run -at a speed of 150 r.p.m., it would then follow that its delivery of air -would amount to: 150-30 = 120. Its volumetric efficiency would be 120 -÷ 150 = 80 per cent. The above figures must not be relied upon, as I -give them simply by way of illustration. - -About a year ago I had the pleasure of examining the tabulated results -of some extensive experiments in this direction, made by one of the -blower companies. I believe they carried their experiments up to 10 lb. -pressure, and I regret that I have not the figures before me, so that -I could give something definite. I do, however, remember that in the -experimental blower, when running at about 150 r.p.m., the volumetric -efficiency at 2 lb. pressure was about 85 per cent., and that at 3 lb. -pressure the volumetric efficiency was about 81 per cent. - -It is unnecessary in this connection to call attention to the -horse-power efficiency of rotary blowers. This is a matter entirely -by itself, and there is considerable difference of opinion among -engineers as to the relative horse-power efficiency of rotary blowers -and piston blowers. All agree that there is a certain pressure at which -the efficiency of the blower becomes less than the efficiency of the -blowing engine. This I have heard placed all the way from 2 lb. up to 6 -lb. - -At the smelting plant of the Tennessee Copper Company we have lately -installed blast-furnace piston-blowing engines; the steam cylinders -are of the Corliss type and are 13 and 24 in. by 42 in.; the blowing -cylinders are two in number, each 57 × 42 in.; the air valves are all -Corliss in type. These blowing engines are designed to operate at a -maximum air pressure of 2½ lb. per square inch. - -At the Santa Fe Gold and Copper Mining Company’s smelter we have -recently installed a No. 8 blower directly coupled to a 14 × 32 in. -Corliss engine. This blower has been in use about five months and is -giving very good results against the comparatively low pressure of 12 -oz., or ¾ lb. - -During the coming summer it is my intention to make careful volumetric -and horse-power tests on these two types of machines under similar -conditions of air pressure, and to publish the results; but in the -meantime I wish to correct the error that a rotary blower of the -impeller type is not a practicable machine at pressure over 5 lb. - - - - - BLOWERS AND BLOWING ENGINES FOR LEAD AND COPPER SMELTING - - BY HIRAM W. HIXON - - (July 20, 1901) - - -In the _Engineering and Mining Journal_ for July 6th I note the -discussion over the relative merits of blowers and blowing engines for -lead and copper smelting, and wish to state that, irrespective of the -work to be done, the blast pressure will depend entirely on the charge -burden in any kind of blast-furnace work, and that the charge burden -governs the reducing action of the furnace altogether. Along these -lines the iron industry has raised the charge burden up to 100 ft. to -secure the full benefit of the reducing action of the carbon monoxide -on the ore. - -In direct opposition to this we have what is known as pyritic smelting, -wherein the charge burden governs the grade of the matte produced to -such an extent that if a charge run with 4 to 6 ft. of burden above the -tuyeres, producing 40 per cent. matte, is changed to a charge burden of -10 or 12 ft., the grade of the matte will decrease from 40 per cent. to -probably less than 20 per cent. I can state this as a fact from recent -experience in operating a blast furnace on heap-roasted ores under the -conditions named, with the result as above stated. - -I was exceedingly skeptical about pyritic smelting as advocated by -some of your correspondents, and still continue to be so; but on -making inquiries from some of my co-workers in this line, Mr. Sticht -of Tasmania, and Mr. Nutting of Bingham, Utah, I have arrived at the -following conclusion, to which some may take exception: That pyritic -smelting without fuel, or with less than 5 per cent., with hot -blast, is practically impossible; that smelting raw ore with a low -charge burden to avoid the reducing action of the carbon monoxide, -thereby securing oxidation of the iron and sulphur, is possible and -practicable, under favorable conditions; and that a large portion of -the sulphur is burned off, and the iron, without reducing action, -goes into the slag in combination with silica. These results can be -obtained with cold blast. - -A blowing engine would certainly be much out of place for operating -copper-matting furnaces run with the evident intention of oxidizing -sulphur and iron and securing as high a grade of matte as possible, -for the reason that to do this it is necessary to run a low charge -burden, and with a low charge burden a high pressure of blast cannot -be maintained. With a 4 to 6 ft. charge burden the blast pressure at -Victoria Mines at present is 3 oz., produced by a No. 6 Green blower -run at 120 r.p.m.; and a blowing engine, delivering the same amount -of air, would certainly not give more pressure. Under the conditions -which we have, a fan would be more effective than a pressure blower, -and a blowing engine entirely out of the question as far as economy is -concerned. - -I installed blowing engines at the East Helena for lead smelting where -the charge burden was 21 ft. and the blast pressure at times went up -as high as 48 oz. Under these conditions the blowing engines gave -satisfaction, but I am of the opinion that the same amount of blast -could have been obtained under that pressure with less horse-power by -the best type of rotary blower. I do not believe that the field of -the blowing engine properly exists below 5 lb., and if this pressure -cannot be obtained by charge-burden conditions, their installation is a -mistake. - -I wish to add the very evident fact that varying the grade of the matte -by feeding the furnace at different hights varies the slag composition -as to its silica and iron contents and makes the feeder the real -metallurgist. The reducing action in the furnace is effected almost -entirely by the gases, and when these are allowed to go to waste, -reduction ceases. - - - - - BLOWING ENGINES AND ROTARY BLOWERS—HOT BLAST FOR PYRITIC SMELTING - - BY S. E. BRETHERTON - - (August 24, 1901) - - -I have just read in the _Engineering and Mining Journal_ of July 20th -an interesting letter written by Hiram W. Hixon in regard to blowing -engines versus the rotary blowers, and also the use of cold blast for -pyritic smelting. - -The controversy, which I unintentionally started in my letter in -the _Engineering and Mining Journal_ of April 13th last, about the -advantages of using either blowers or blowing engines for blast -furnaces, does not particularly interest me, with the exception that I -have about decided, in my own mind, to use blowing engines where there -is much back pressure, and the ordinary up-to-date blower for pyritic -or matte smelting where much back pressure should not exist. I fully -appreciate the fact that so-called pyritic smelting can be done to a -limited extent, even with cold blast. Theoretically, enough oxygen -can be sent into the blast furnace, contained in the cold blast, to -oxidize both the fuel and the sulphur in an ordinary sulphide charge, -but I have not yet learned where a high concentration is being made -with unroasted ore and cold blast. I experimented on these lines at -different times for three years, during 1896, 1897, and 1898, making -a fair concentration with refractory ores, most of which had been -roasted. I was myself interested in the profits and as anxious as any -one for economy. We tried, for fuel in the blast furnace, coke alone, -coke and lignite coal, lignite coal alone, lignite coal and dry wood, -coal and green wood, and then coke and green wood, all under different -hights of ore burden in the furnace. - -A description of these experiments would, no doubt, be tiresome to your -readers, but I wish to state that the furnace was frozen up several -times on account of using too little fuel, when the cold blast would -gradually drive nearly all the heat to the top of the furnace, the -crucible and between the tuyeres becoming so badly crusted that the -furnace had to be cleaned out and blown in again, unless I was called -in time to save it by changing the charge and increasing the fuel. We -were making high-grade matte under contract, high concentration and -small matte fall, which would, no doubt, aggravate matters. - -After the introduction of hot blast, heated up to between 200 and 300 -deg. F., we made the same grade of matte from the same character of -ore, with the exception that we then smelted without roasting, and -reduced the percentage of fuel consumption, increased the capacity of -the furnace, and almost entirely obviated the trouble of cold crucibles -and hot tops. I write the above facts, as they speak for themselves. - -I nearly agree with Mr. Hixon, and do not think it practical to smelt -with much less than 5 per cent. coke continuously; but there is a -great saving between the amount of coke used with a moderately heated -blast and cold blast. Regardless of either hot or cold blast, however, -the fuel consumption depends very much on the character of the ore -to be smelted, the amount of matte-fall and grade of matte made. It -is not always advisable or necessary to use hot blast for a matting -furnace; that is, where the supply of sulphur is limited. It may then -be necessary to use as much fuel in the blast furnace to prevent the -sulphur from oxidizing as will be sufficient to furnish the heat for -smelting. Such conditions existed at Silver City, N. M. , at times, -after our surplus supply of iron and zinc sulphide concentrates was -used. I understand that they are now short of sulphur there, on account -of getting a surplus amount of oxidized copper ore, and are only -utilizing what little heat the slag gives them, without the addition of -any fuel on top of the forehearth. - -Before closing this, which I intended to have been brief, I wish to -call your attention to a little experience we had with alumina in the -matting furnace at Silverton, Col., where I was acting as consulting -metallurgist. The ore we had to smelt contained, on an average, about -20 per cent. Al₂O₃, 30 per cent. SiO₂, with 18 per cent. Fe in -the form of an iron pyrite, and no other iron was available except some -iron sulphide concentrates containing a small percentage of zinc and -lead. - -The question naturally arose, could we oxidize and force sufficient -of the iron into the slag, and where should we class the alumina, as -a base or an acid? My experience in lead smelting led me to believe -that Al₂O₃ could only be classed as an acid in the ordinary -lead furnace, and that it would be useless to class it otherwise in -a shallow matting furnace; and E. W. Walter, the superintendent and -metallurgist in charge, agreed with me. - -We then decided to make a bisilicate slag, classing the alumina as -silica, and we obtained fairly satisfactory results. The slag made -was very clean, but treacherous, which was attributed to two reasons: -First, that it required more heat to keep the alumina slag liquid -enough to flow than it does a nearly straight silica slag; and, second, -that we were running so close to the formula of a bisilicate and -aluminate slag (about 31½ per cent. SiO₂, 27 per cent. Fe, 20 per -cent. CaO, and 18 per cent. Al₂O₃, or 49½ per cent. acid) that a -few charges thrown into the furnace containing more silica or alumina -than usual would thicken the slag so that it would then require some -extra coke and flux to save the furnace. At times the combined SiO₂ -and Al₂O₃ did reach 55 and 56 per cent. in the slag, which did -not freeze up the furnace, but caused us trouble. - - - - - PART IX - - LEAD REFINING - - - - - THE REFINING OF LEAD BULLION[49] - - BY F. L. PIDDINGTON - - (October 3, 1903) - - -In presenting this account of the Parkes process of desilverizing and -refining lead bullion no originality is claimed, but I hope that a -description of the process as carried out at the works of the Smelting -Company of Australia may be of service. - -_Introductory._—The Parkes process may be conveniently summarized as -follows: - -1. Softening of the base bullion to remove copper, antimony, etc. - -2. Removal of precious metals from the softened bullion by means of -zinc. - -3. Refining the desilverized lead. - -4. Liquation of gold and silver crusts obtained from operation No. 2. - -5. Retorting the liquated alloy to drive off zinc. - -6. Concentrating and refining bullion from No. 5. - -_Softening._—This is done in reverberatory furnaces. In large works two -furnaces are used, copper, antimony, and arsenic being removed in the -first and antimony in the second. The size of the furnaces is naturally -governed by the quantity to be treated. In these works (refining about -200 tons weekly) a double set of 15-ton furnaces were at work. The -sides and ends of these furnaces are protected by a jacket with a 2-in. -water space, the jacket extending some 3 in. above the charge level and -6 in. to 9 in. below it. The furnace is built into a wrought-iron pan, -and if the brickwork is well laid into the pan there need be no fear of -lead breaking through below the jacket. - -The bars of bullion (containing, as a rule, 2 to 3 per cent. of -impurities) are placed in the furnace carefully, to avoid injuring the -hearth, and melted down slowly. The copper dross separates out and -floats on top of the charge, which is stirred frequently to expose -fresh surfaces. If the furnace is overheated some dross is melted into -the lead again and will not separate out until the charge is cooled -back. However carefully the work is done some copper remains with the -lead, and its effects are to be seen in the later stages. The dross is -skimmed into a slag pot with a hole bored in it some 4 in. from the -bottom; any lead drained from the pot is returned to the charge. The -copper dross is either sent back to the blast furnace direct or may -be first liquated. By the latter method some 30 per cent. of the lead -contents of the dross is recovered in the refinery. - -Base bullion made at a customer’s smelter will often vary greatly in -composition, and it is, therefore, difficult to give any hard and fast -figures as to percentage of metals in the dross. As a rule our dross -showed 65 to 70 per cent. lead, copper 2 to 9 per cent. (average 4 per -cent.), gold and silver values varying with the grade of the original -bullion, though it was difficult to detect any definite relation -between bullion and dross. It was, however, noticed that gold and -silver values increased with the percentage of copper. - -Immediately the copper dross is skimmed off the heat is raised -considerably, and very soon a tin (and arsenic, if present) skimming -appears. It is quite “dry” and may be removed in an hour or so. It is a -very small skimming, and the tin, not being worth saving, is put with -the copper dross. - -The temperature is now raised again and antimony soon shows in black, -boiling, oily drops, gathering in time into a sheet covering the -surface of the lead. When the skimming is about ½-inch thick, slaked -lime, ashes, or fine coal is thrown on and stirred in. The dross soon -thickens up and may be skimmed off easily. This operation is repeated -until all antimony is eliminated. Constant stirring of the charge is -necessary. The addition of litharge greatly facilitates the removal -of antimony; either steam or air may be blown on the surface of the -metal to hasten oxidation, though they have anything but a beneficial -effect on the furnace lining. From time to time samples of the dross -are taken in a small ladle, and after setting hard the sample is broken -in two. A black vitreous appearance indicates plenty of antimony yet -in the charge. Later samples will look less black, until finally a few -yellowish streaks are seen, being the first appearance of litharge. -When all antimony is out the fracture of a sample should be quite -yellow and the grain of the litharge long, a short grain indicating -impurities still present, in which case another skimming is necessary. -The analysis of a representative sample of antimony dross was as -follows: - - PbO = 78.11 per cent. - Sb₂O₄ = 8.75 ” ” - As₂O₃ = 2.18 ” ” - CuO = 0.36 ” ” - CaO = 1.10 per cent. - Fe₂O₃ = 0.42 ” ” - Al₂O₃ = 0.87 ” ” - Insol. = 4.10 ” ” - -Antimony dross is usually kept separate and worked up from time to -time, yielding hard antimonial lead, used for type metal, Britannia -metal, etc. - -_Desilverization._—The softening being completed the charge is tapped -and run to a kettle or pan of cast iron or steel, holding, when -conveniently full, some 12 or 13 tons. The lead falling into the -kettle forms a considerable amount of dross, which is skimmed off and -returned to the softening furnace. By cooling down the charge until -it nearly “freezes” an additional copper skimming is obtained, which -also is returned to the softener. The kettle is now heated up to the -melting point of zinc, and the zinc charge, determined by the gold -and silver contents of the kettle, is added and melted. The charge -is stirred, either by hand or steam, for about an hour, after which -the kettle is allowed to cool down for some three hours and the first -zinc crust taken off. When the charge is skimmed clean a sample of the -bullion is taken for assay, and while this is being done the kettle is -heated again for the second zinc charge, which is worked in the same -way as the first; sometimes a third addition of zinc is necessary. The -resulting crusts are kept separate, the second and third being added -to the next charge as “returns,” allowing 3 lb. of zinc in returns as -equal to 1 lb. of fresh zinc. An alternative method is to take out gold -and silver in separate crusts, in which case the quantity of zinc first -added is calculated on the gold contents of the kettle only. The method -of working is the same, though subsequent treatment may differ in that -the gold crusts are cupeled direct. - -As to the quantity of zinc required: - -1. Extracting the gold with as little silver as possible, the following -figures were obtained: - - Total Gold— Au. - In kettle 300 oz. │ 1 lb. zinc takes out 1.00 oz. - ” ” 200 ” │ ” ” ” ” 1.00 ” - ” ” 150 ” │ ” ” ” ” 0.79 ” - ” ” 100 ” │ ” ” ” ” 0.59 ” - ” ” 60 ” │ ” ” ” ” 0.45 ” - -2. Silver zinking gave the following general results with 11-ton -charges: - - Total Silver— - In kettle 1,450 oz. │ 1 lb. zinc takes out 5.6 oz. - ” ” 1,200 ” │ ” ” ” ” 4.1 ” - ” ” 930 ” │ ” ” ” ” 3.8 ” - ” ” 755 ” │ ” ” ” ” 3.5 ” - ” ” 616 ” │ ” ” ” ” 3.4 ” - ” ” 460 ” │ ” ” ” ” 2.6 ” - -3. Extracting gold and silver together: - - ───────────────────────────┬────────────────────── - TOTAL CONTENTS OF KETTLE │ 1 LB. ZINC TAKES OUT - AU. OZ. │ AG. OZ. │ AU. OZ. │ AG. OZ. - ────────────┼──────────────┼─────────────┼──────── - 494 │ 3,110 │ 0.59 │ 3.60 - 443 │ 1,883 │ 0.64 │ 2.80 - 330 │ 2,417 │ 0.45 │ 3.34 - 204 │ 1,638 │ 0.36 │ 2.86 - 143 │ 1,330 │ 0.28 │ 2.65 - 123 │ 1,320 │ 0.23 │ 2.54 - ────────────┴──────────────┴─────────────┴──────── - -It will be noticed that in each case the richer the bullion the greater -the extractive power of zinc. Experiments made on charges of rich -bullion showed that the large amount of zinc called for by the table in -use was unnecessary, and 250 lb. was fixed on as the first addition of -zinc. On this basis an average of 237 charges gave results as follows: - - ───────────────────┬───────────┬────────────────────── - TOTAL CONTENTS │ ZINC USED │ 1 LB. ZINC TAKES OUT - AU. OZ. │ AG. OZ. │ LBS. │ AU. OZ. │ AG. OZ. - ────────┼──────────┼───────────┼───────────┼────────── - 520 │ 1,186 │ 507.5 │ 1.27 │ 2.91 - ────────┴──────────┴───────────┴───────────┴────────── - -The zinc used was that necessary to clean the kettle, added as follows: -1st, 250 lb.; 2d (average), 127 lb.; 3d (average), 57 lb. In 112 cases -no third addition was required. From these figures it appears that in -the earlier work the zinc was by no means saturated. - -_Refining the Lead._—Gold and silver being removed, the lead is -siphoned off into the refining kettle and the fire made up. In about -four hours the lead will be red hot, and when hot enough to burn zinc, -dry steam, delivered by a ¾ in. pipe reaching nearly to the bottom of -the kettle, is turned on. The charge is stirred from time to time and -wood is fed on the top to assist dezinking and prevent the formation -of too much litharge. In three to four hours the lead will be soft and -practically free from zinc. When test strips show the lead to be quite -soft and clean, the kettle is cooled down and the scum of lead and -zinc oxides skimmed off. In an hour or so the lead will be cool enough -for molding; the bar should have a yellow luster on the face when set; -if the lead is too cold it will be white, if too hot a deep blue. The -refining kettles are subjected to severe strain during the steaming -process, and hence their life is uncertain—an average would be about 60 -charges; the zinking kettles, on the other hand, last very much longer. -Good steel kettles (if they can be obtained) are preferable to cast -iron. - -_Treatment of Zinc Crusts._—Having disposed of the lead, let us -return now to the zinc crusts. These are first liquated in a small -reverberatory furnace, the hearth of which is formed of a cast-iron -plate (the edges of the long sides being turned up some 4 in.) laid on -brasque filling, with a fall from bridge to flue of ¾ in. per foot and -also sloping from sides to center. The operation is conducted at a low -temperature and the charge is turned over at intervals, the liquated -lead running out into a small separately fired kettle. This lead rarely -contains more than a few ounces of silver per ton; it is baled into -bars, and returned to the zinking kettles or worked up in a separate -charge. In two to three hours the crust is as “dry” as it is advisable -to make it, and the liquated alloy is raked out over a slanting -perforated plate to break it up and goes to the retort bin. - -_Retorting the Alloy._—This is carried on in Faber du Faur tilting -furnaces—simply a cast-iron box swinging on trunnions and lined with -firebrick. Battersea retorts (class 409) holding 560 lb. each are -used; their average life is about 30 charges. The retorts are charged -hot, a small shovel of coal being added with the alloy. The condenser -is now put in place and luted on; it is made of ⅛ in. iron bent to -form a cylinder 12 in. in diameter, open at one end; it is lined with -a mixture of lime, clay and cement. It has three holes, one on the -upper side close to the furnace and through which a rod can be passed -into the retort, a vent-hole on the upper side away from the furnace, -and a tap-hole on the bottom for condensed zinc. In an hour or so the -flame from the vent-hole should be green, showing that distillation has -begun. When condensation ceases (shown by the flame) the condenser is -removed and the bullion skimmed and poured into bars for the cupel. The -products of retorting are bullion, zinc, zinc powder and dross. Bullion -goes to the cupel, zinc is used again in the desilverizing kettles, -powder is sieved to take out scraps of zinc and returned to the blast -furnace, or it may be, and sometimes is, used as a precipitating -agent in cyanide works; dross is either sweated down in a cupel with -lead and litharge, together with outside material such as zinc gold -slimes from cyanide works, jeweler’s sweep, mint sweep, etc., or in -the softening furnace after the antimony has been taken off. In either -case the resulting slag goes back to the blast furnace. The total -weight of alloy treated is approximately 7 per cent. of the original -base bullion. The zinc recovered is about 60 per cent. of that used -in desilverizing. The most important source of temporary loss is the -retort dross (consisting of lead-zinc-copper alloy with carbon, silica -and other impurities), and it is here that the necessity of removing -copper in the softening process is seen, since any copper comes out -with the zinc crusts and goes on to the retorts, where it enters the -dross, carrying gold and silver with it. If much copper is present the -dross may contain more gold and silver than the retort bullion itself. -In this connection I remember an occasion on which some retort dross -yielded gold and silver to the extent of over 800 and 3000 oz. per ton -respectively. - -_Cupellation._—Retort bullion is first concentrated (together with -bullion resulting from dross treatment) to 50 to 60 per cent. gold and -silver in a water-jacketed cupel. The side lining is protected by an -inch water-pipe imbedded in the lining at the litharge level or by a -water-jacket, the inner face of which is of copper; the cupel has also -a water-jacketed breast so that the front is not cut down. The cupel -lining may be composed of limestone, cement, fire-clay and magnesite -in various proportions, but a simple lining of sand and cement was -found quite satisfactory. When the bullion is concentrated up to 50 to -60 per cent. gold and silver, it is baled out and transferred to the -finishing cupel, where it is run up to about 0.995 fine; it is then -ready either for the melting-pot or parting plant. The refining test, -by the way, is not water-cooled. - -Re-melting is done in 200-oz. plumbago crucibles and presents no -special features. In the case of doré bullion low in gold, “sprouting” -of the silver is guarded against by placing a piece of wood or charcoal -on the surface of the metal before pouring, and any slag is kept back. -The quantity of slag formed is, of course, very small, so that the bars -do not require much cleaning. - -The parting plant was not in operation in my time, and I am therefore -unable to go into details. The process arranged for was briefly as -follows: Solution of the doré bullion in H₂SO₄; crystallization -of silver as monosulphate by dilution and cooling; decomposition of -silver sulphate by ferrous sulphate solution giving metallic silver and -ferric sulphate, which is reduced to the ferrous salt by contact with -scrap iron. The gold and silver are washed thoroughly with hot water -and cast into bars. - -In conclusion, some variations in practice may be noted. The use of -two furnaces in the softening process has already been mentioned; by -this means the drossing and softening are more perfect and subsequent -operations thereby facilitated; further, the furnaces, being kept at -a more equable temperature, are less subject to wear and tear. Zinc -crusts are sometimes skimmed direct into an alloy press in which -the excess of lead is squeezed out while still molten; liquation is -then unnecessary. Refining of the lead may be effected by a simple -scorification in a reverberatory, the soft lead being run into a kettle -from which it is molded into market bars. - - - - - THE ELECTROLYTIC REFINING OF BASE LEAD BULLION - - BY TITUS ULKE - - (October 11, 1902) - - -Important changes in lead-refining practice are bound to follow, in my -opinion, the late demonstration on a large scale of the low working -cost and high efficiency of Betts’ electrolytic process of refining -lead bullion. It was my good fortune recently to see this highly -interesting process in operation at Trail, British Columbia, through -the kindness of the inventor, A. G. Betts, and Messrs. Labarthe and -Aldridge, of the Trail works. - -A plant of about 10 tons daily capacity, which probably cost about -$25,000, although it could be duplicated for perhaps $15,000 at the -present time, was installed near the Trail smelting works. It has been -in operation for about ten months, I am informed, with signal success, -and the erection of a larger plant, of approximately 30 tons capacity -and provided with improved handling facilities, is now completed. - -The depositing-room contains 20 tanks, built of wood, lined with tar, -and approximately of the size of copper-refining tanks. Underneath the -tank-room floor is a basement permitting inspection of the tank bottoms -for possible leakage and removal of the solution and slime. A suction -pump is employed in lifting the electrolyte from the receiving tank and -circulating the solution. In nearly every respect the arrangement of -the plant and its equipment is strikingly like that of a modern copper -refinery. - -The great success of the process is primarily based upon Betts’ -discovery of the easy solubility of lead in an acid solution of lead -fluosilicate, which possesses both stability under electrolysis and -high conductivity, and from which exceptionally pure lead may be -deposited with impure anodes at a very low cost. With such a solution, -there is no polarization from formation of lead peroxide on the anode, -no evaporation of constituents except water, and no danger in its -handling. It is cheaply obtained by diluting hydrofluoric acid of -35 per cent. HF, which is quoted in New York at 3c. per pound, with -an equal volume of water and saturating it with pulverized quartz -according to the equation: - - SiO₂ + 6HF = HSiF₆ + 2H₂O. - -According to Mr. Betts, an acid of 20 to 22 per cent. will come -to about $1 per cu. ft., or to $1.25 when the solution has been -standardized with 6 lb. of lead. One per cent. of lead will neutralize -0.7 per cent. H₂SiF₆. The electrolyte employed at the time of my -inspection of the works contained, I believe, 8 per cent. lead and 11 -per cent. excess of fluosilicic acid. - -The anodes consist of the lead bullion to be refined, cast into plates -about 2 in. thick and approximately of the same size as ordinary -two-lugged copper anodes. Before being placed in position in the tanks, -they are straightened by hammering over a mold and their lugs squared. -No anode sacks are employed as in the old Keith process. - -The cathode sheets which receive the regular lead deposits are thin -lead plates obtained by electrodeposition upon and stripping from -special cathodes of sheet steel. The latter are prepared for use by -cleaning, flashing with copper, lightly lead-plating in the tanks, and -greasing with a benzine solution of paraffin, dried on, from which the -deposited lead is easily stripped. - -The anodes and cathodes are separated by a space of 1½ to 2 in. in the -tank and are electrically connected in multiple, the tanks being in -series circuit. The fall in potential between tanks is only about 0.2 -of a volt, which remarkably low voltage is due to the high conducting -power of the electrolyte and to some extent to the system of contacts -used. These contacts are small wells of mercury in the bus-bars, large -enough to accommodate copper pins soldered to the iron cathodes or -clamped to the anodes. Only a small amount of mercury is required. - -Current strengths of from 10 to 25 amperes per sq. ft. have been used, -but at Trail 14 amperes have given the most satisfactory results as -regards economy of working and the physical and chemical properties of -the refined metal produced. - -A current of 1 ampere deposits 3.88 grams of lead per hour, or -transports 3¼ times as much lead, in this case, as copper with an -ordinary copper-refining solution. A little over 1000 kg., or 2240 -lb., requires about 260,000 ampere hours. At 10 amperes per sq. ft. the -cathode (or anode) area should be about 1080 sq. ft. per ton of daily -output. Taking a layer of electrolyte 1.5 in. thick, 135 cu. ft. will -be found to be the amount between the electrodes, and 175 cu. ft. may -be taken as the total quantity of solution necessary, according to Mr. -Betts’ estimate. The inventor states that he has worked continuously -and successfully with a drop of potential of only 0.175 volt per tank, -and that therefore 0.25 volt should be an ample allowance in regular -refining. Quoting Mr. Betts; “260,000 ampere hours at 0.25 volt works -out to 87 electrical h.p. hours of 100 h.p. hours at the engine shaft, -in round numbers. Estimating that 1 h.p. hour requires the burning of -1.5 lb. of coal, and allowing say 60 lb. for casting the anodes and -refined lead, each ton of lead refined requires the burning of 210 lb. -of fuel.” With coal at $6 per ton the total amount of fuel consumed, -therefore, should not cost over 60c., which is far below the cost of -fire-refining base lead bullion, as we know. - -In the Betts electrolytic process, practically all the impurities -in the base bullion remain as a more or less adherent coating on -the anode, and only the zinc, iron, cobalt and nickel present go -into solution. The anode residue consists practically of all the -copper, antimony, bismuth, arsenic, silver and gold contained in the -bullion, and very nearly 10 per cent. of its weight in lead. Having -the analysis of any bullion, it is easy to calculate with these data -the composition of the anode residue and the rate of pollution of the -electrolyte. Allowing 175 cu. ft. of electrolyte per ton of daily -output, it will be found that in the course of a year these impurities -will have accumulated to the extent of a very few per cent. Estimating -that the electrolyte will have to be purified once a year, the amount -to be purified daily is less than 1 cu. ft. for each ton of output. -The amount of lead not immediately recovered in pure form is about -0.3 per cent., most of which is finally recovered. As compared with -the ordinary fire-refined lead, the electrolytically refined lead is -much purer and contains only mere traces of bismuth, when bismuthy -base bullion is treated. Furthermore, the present loss of silver in -fire refining, amounting, it is claimed, to about 1½ per cent. of the -silver present, and covered by the ordinary loss in assay, is to a -large extent avoided, as the silver in the electrolytic process is -concentrated in the anode residue with a very small loss, and the loss -of silver in refining the slimes is much less than in treating the -zinc crusts and refining the silver residue after distillation. The -silver slimes obtained at Trail, averaging about 8000 oz. of gold and -silver per ton, are now treated at the Seattle Smelting and Refining -Works. There the slimes are boiled with concentrated sulphuric acid and -steam, allowing free access of air, which removes the greater part of -the copper. The washed residue is then dried in pans over steam coils, -and melted down in a magnesia brick-lined reverberatory, provided -with blast tuyeres, and refined. In this reverberatory furnace the -remainder of the copper left in the slimes after boiling is removed by -the addition of niter as a flux, and the antimony with soda. The doré -bars finally obtained are parted in the usual way with sulphuric acid, -making silver 0.999 fine and gold bars at least 0.992 fine. - -Mr. Betts treated 2000 grams of bullion, analyzing 98.76 per cent. Pb, -0.50 Ag, 0.31 Cu, and 0.43 Sb with a current of 25 amp. per square -foot in an experimental way, and obtained products of the following -composition: - -Refined Lead: 99.9971 per cent. Pb, 0.0003 Ag, 0.0007 Cu, and 0.0019 Sb. - -Anode Residue: 9.0 per cent. Pb, 36.4 Ag, 25.1 Cu, and 2.95 Sb. - -Four hundred and fifty pounds of bullion from the Compania Metalurgica -Mexicana, analyzing 0.75 per cent. Cu, 1.22 Bi, 0.94 As, 0.68 Sb, and -assaying 358.9 oz. Ag and 1.71 oz. Au per ton, were refined with a -current of 10 amp. per square foot, and gave a refined lead of the -following analysis: 0.00027 per cent. Cu, 0.0037 Bi, 0.0025 As, 0 Sb, -0.0010 Ag, 0.0022 Fe, 0.0018 Zn and Pb (by difference) 99.9861 per cent. - -Although the present method for recovering the precious metals and -by-products from the anode residue leaves much room for improvement, -the use of the Betts process may be recommended to our lead refiners, -because it is a more economical and efficient method than the -fire-refining process now in common use. I will state my belief, in -conclusion, that the present development of electrolytic lead refining -signalizes as great an advance over zinc desilverization and the fire -methods of refining lead as electrolytic copper refining does over the -old Welsh method of refining that metal. - - - - - ELECTROLYTIC LEAD-REFINING[50] - - BY ANSON G. BETTS - - -A solution of lead fluosilicate, containing an excess of fluosilicic -acid, has been found to work very satisfactorily as an electrolyte -for refining lead. It conducts the current well, is easily handled -and stored, non-volatile and stable under electrolysis, may be made -to contain a considerable amount of dissolved lead, and is easily -prepared from inexpensive materials. It possesses, however, in common -with other lead electrolytes, the defect of yielding a deposit of lead -lacking in solidity, which grows in crystalline branches toward the -anodes, causing short circuits. But if a reducing action (practically -accomplished by the addition of gelatine or glue) be given to the -solution, a perfectly solid and dense deposit is obtained, having very -nearly the same structure as electrolytically deposited copper, and a -specific gravity of about 11.36, which is that of cast lead. - -Lead fluosilicate may be crystallized in very soluble brilliant -crystals, resembling those of lead nitrate and containing -four molecules of water of crystallization, with the formula -PbSiF₆,4H₂O. This salt dissolves at 15 deg. C. in 28 per cent. of -its weight of water, making a syrupy solution of 2.38 sp. gr. Heated -to 60 deg. C., it melts in its water of crystallization. A neutral -solution of lead fluosilicate is partially decomposed on heating, with -the formation of a basic insoluble salt and free fluosilicic acid, -which keeps the rest of the salt in solution. This decomposition ends -when the solution contains perhaps 2 per cent. of free acid; and the -solution may then be evaporated without further decomposition. The -solutions desired for refining are not liable to this decomposition, -since they contain much more than 2 per cent. of free acid. The -electrical conductivity depends mainly on the acidity of the solution. - -My first experiments were carried out without the addition of gelatine -to the fluosilicate solution. The lead deposit consisted of more or -less separate crystals that grew toward the anode, and, finally, caused -short circuits. The cathodes, which were sheet-iron plates, lead-plated -and paraffined, had to be removed periodically from the tanks and -passed through rolls, to pack down the lead. When gelatine has been -added in small quantities, the density of the lead is greater than can -be produced by rolling the crystalline deposit, unless great pressure -is used. - -The Canadian Smelting Works, Trail, B. C. , have installed a refinery, -making use of this process. There are 28 refining-tanks, each 86 in. -long, 30 in. wide and 42 in. deep, and each receiving 22 anodes of -lead bullion with an area of 26 by 33 in. exposed to the electrolyte -on each side, and 23 cathodes of sheet lead, about 1/16 in. thick, -prepared by deposition on lead-plated and paraffined iron cathodes. The -cathodes are suspended from 0.5 by 1 in. copper bars, resting crosswise -on the sides of the tanks. The experiment has been thoroughly tried of -using iron sheets to receive a deposit thicker than 1/16 in.; that is, -suitable for direct melting without the necessity of increasing its -weight by further deposition as an independent cathode; but the iron -sheets are expensive, and are slowly pitted by the action of the acid -solution; and the lead deposits thus obtained are much less smooth and -pure than those on lead sheets. - -The smoothness and the purity of the deposited lead are proportional. -Most of the impurity seems to be introduced mechanically through the -attachment of floating particles of slime to irregularities on the -cathodes. The effect of roughness is cumulative; it is often observed -that particles of slime attract an undue amount of current, resulting -in the lumps seen in the cathodes. Samples taken at the same time -showed from 1 to 2.5 oz. silver per ton in rough pieces from the iron -cathodes, 0.25 oz. as an average for the lead-sheet cathodes, and only -0.04 oz. in samples selected for their smoothness. The variation in -the amount of silver (which is determined frequently) in the samples -of refined lead is attributed not to the greater or less turbidity of -the electrolyte at different times, but to the employment of new men in -the refinery, who require some experience before they remove cathodes -without detaching some slime from the neighboring anodes. - -Each tank is capable of yielding, with a current of 4000 amperes, -750 lb. of refined lead per day. The voltage required to pass this -current was higher than expected, as explained below; and for this -reason, and also because the losses of solution were very heavy until -proper apparatus was put in to wash thoroughly the large volume of -slime produced (resulting in a weakened electrolyte), the current used -has probably averaged about 3000 amperes. The short circuits were -also troublesome, though this difficulty has been greatly reduced by -frequent inspection and careful placing of the electrodes. At one time, -the solution in use had the following composition in grams per 100 -c.c.: Pb, 6.07; Sb, 0.0192; Fe, 0.2490; SiF₆, 6.93, and As, a trace. -The current passing was 2800 amperes, with an average of about 0.44 -volts per tank, including bus-bars and contacts. It is not known what -was the loss of efficiency on that date, due to short circuits; and -it is, therefore, impossible to say what resistance this electrolyte -constituted. - -Hydrofluoric acid of 35 per cent., used as a starting material for the -preparation of the electrolyte, is run by gravity through a series of -tanks for conversion into lead fluosilicate. In the top tank is a layer -of quartz 2 ft. thick, in passing through which the hydrofluoric acid -dissolves silica, forming fluosilicic acid. White lead (lead carbonate) -in the required quantity is added in the next tank, where it dissolves -readily and completely with effervescence. All sulphuric acid and any -hydrofluoric acid that may not have reacted with silica settle out -in combination with lead as lead sulphate and lead fluoride. Lead -fluosilicate is one of the most soluble of salts; so there is never -any danger of its crystallizing out at any degree of concentration -possible under this method. The lead solution is then filtered and run -by gravity into the refining-tanks. - -The solution originally used at Trail contained about 6 per cent. Pb -and 15 per cent. SiF₆. - -The electrical resistance in the tanks was found to be greater than -had been calculated for the same solution, plus an allowance for -loss of voltage in the contacts and conductors. This is partly, at -least, due to the resistance to free motion of the electrolyte, in -the neighborhood of the anode, offered by a layer of slime which may -be anything up to ½ in. thick. During electrolysis, the SiF₆ ions -travel toward the anodes, and there combine with lead. The lead and -hydrogen travel in the opposite direction and out of the slime; but -there are comparatively few lead ions present, so that the solution -in the neighborhood of the anodes must increase in concentration and -tend to become neutral. This greater concentration causes an e.m.f. of -polarization to act against the e.m.f. of the dynamo. This amounted -to about 0.02 volt for each tank. The greater effect comes from the -greater resistance of the neutral solution with which the slime is -saturated. There is, consequently, an advantage in working with rather -thin anodes, when the bullion is impure enough to leave slime sticking -to the plates. A compensating advantage is found in the increased ease -of removing the slime with the anodes, and wiping it off the scrap in -special tanks, instead of emptying the tanks and cleaning out, as is -done in copper refineries. - -It is very necessary to have adequate apparatus for washing solution -out of the slime. The filter first used consisted of a supported -filtering cloth with suction underneath. It was very difficult to -get this to do satisfactory work by reason of the large amount of -fluosilicate to be washed out with only a limited amount of water. -At the present time the slime is first stirred up with the ordinary -electrolyte several times, and allowed to settle, before starting to -wash with water at all. The Trail plant produces daily 8 or 10 cu. ft. -of anode residue, of which over 90 per cent. by volume is solution. -The evaporation from the total tank surface of something like 400 sq. -ft. is only about 15 cu. ft. daily; so that only a limited amount of -wash-water is to be used—namely, enough to replace the evaporated -water, plus the volume of the slime taken out. - -The tanks are made of 2 in. cedar, bolted together and thoroughly -painted with rubber paint. Any leakage is caught underneath on sloping -boards. Solution is circulated from one tank to another by gravity, and -is pumped from the lowest to the highest by means of a wooden pump. The -22 anodes in each tank together weigh about 3 tons, and dissolve in -from 8 to 10 days, two sets of cathodes usually being used with each -set of anodes. While 300 lb. cathodes can be made, the short-circuiting -gets so troublesome with the spacing used that the loss of capacity is -more disadvantageous than the extra work of putting in and taking out -more plates. The lead sheets used for cathodes are made by depositing -about 1/16 in. metal on paraffined steel sheets in four of the tanks, -which are different from the others only in being a little deeper. - -The anodes may contain any or all of the elements, gold, silver, -copper, tin, antimony, arsenic, bismuth, cadmium, zinc, iron, nickel, -cobalt and sulphur. It would be expected that gold, silver, copper, -antimony, arsenic and bismuth, being more electronegative than lead, -would remain in the slime in the metallic state, with, perhaps, tin, -while iron, zinc, nickel and cobalt would dissolve. It appears that tin -stands in the same relation to lead that nickel does to iron, that is, -they have about the same electromotive forces of solution, with the -consequence that they can behave as one metal and dissolve and deposit -together. Iron, contrary to expectation, dissolves only slightly, while -the slime will carry about 1 per cent. of it. It appears from this that -the iron exists in the lead in the form of matte. Arsenic, antimony, -bismuth and copper have electromotive forces of solution more than 0.3 -volt below that of lead. As there is no chance that any particle of -one of these impurities will have an electric potential of 0.3 volt -above that of the lead with which it is in metallic contact, there is -no chance that they will be dissolved by the action of the current. The -same is even more certainly true of silver and gold. The behavior of -bismuth is interesting and satisfactory. It is as completely removed by -this process of refining as antimony is. No other process of refining -lead will remove this objectionable impurity so completely. Tin has -been found in the refined lead to the extent of 0.02 to 0.03 per cent. -This we had no difficulty in removing from the lead by poling before -casting. There is always a certain amount of dross formed in melting -down the cathodes; and the lead oxide of this reacts with the tin in -the lead at a comparatively low temperature. - -The extra amount of dross formed in poling is small, and amounts to -less than 1 per cent. of the lead. The dross carries more antimony and -arsenic than the lead, as well as all the tin. The total amount of -dross formed is about 4 per cent. Table I shows its composition. - -The electrolyte takes up no impurities, except, possibly, a small part -of the iron and zinc. Estimating that the anodes contain 0.01 per -cent. of zinc and soluble iron, and that there are 150 cu. ft. of the -solution in the refinery for every ton of lead turned out daily, in -one year the 150 cu. ft. will have taken up 93 lb. of iron and zinc, -or about 1 per cent. These impurities can accumulate to a much greater -extent than this before their presence will become objectionable. It -is possible to purify the electrolyte in several ways. For example, -the lead can be removed by precipitation with sulphuric acid, and -the fluosilicic acid precipitated with salt as sodium fluosilicate. -By distillation with sulphuric acid the fluosilicic acid could be -recovered, this process, theoretically, requiring but one-third as much -sulphuric acid as the decomposition of fluorspar, in which the fluorine -was originally contained. - -The only danger of lead-poisoning to which the workmen are exposed -occurs in melting the lead and casting it. In this respect the -electrolytic process presents a distinct sanitary advance. - -For the treatment of slime, the only method in general use consists in -suspending the slime in a solution capable of dissolving the impurities -and supplying, by a jet of steam and air forced into the solution, the -air necessary for its reaction with, and solution of, such an inactive -metal as copper. After the impurities have been mostly dissolved, the -slime is filtered off, dried and melted, under such fluxes as soda, to -a doré bullion. - -The amount of power required is calculated thus: Five amperes in 24 -hours make 1 lb. of lead per tank. One ton of lead equals 10,000 -ampere-days, and at 0.35 volts per tank, 3500 watt-days, or 4.7 -electric h.p. days. Allowing 10 per cent. loss of efficiency in the -tanks (we always get less lead than the current which is passing would -indicate), and of 8 per cent. loss in the generator, increases this to -about 5.6 h.p. days, and a further allowance for the electric lights -and other applications gives from 7 to 8 h.p. days as about the amount -per ton of lead. At $30 per year, this item of cost is something like -65c. per ton of lead. So this is an electro-chemical process not -especially favored by water-power. - -The cost of labor is not greater than in the zinc-desilverization -process. A comparison between this process and the Parkes process, on -the assumption that the costs for labor, interest and general expenses -are about equal, shows that about $1 worth of zinc and a considerable -amount of coal and coke have been done away with, at the expense -of power, equal to about 175 h.p. hours, of the average value of -perhaps 65c., and a small amount of coal for melting the lead in the -electrolytic method. - -More important, however, is the greater saving of the metal values by -reason of increased yields of gold, silver, lead, antimony and bismuth, -and the freedom of the refined lead from bismuth. - -Tables II, III, and IV show the composition of bullion, slimes and -refined lead. - -Tables V, VI, VII, and VIII give the results obtained experimentally in -the laboratory on lots of a few pounds up to a few hundred pounds. The -results in Tables VI and VII were given me by the companies for which -the experiments were made. - - -TABLE I.—ANALYSES OF DROSS - -For analyses of the lead from which this dross was taken, see Table II - - ───┬──────┬─────────┬─────────┬─────────┬─────────┬───── - │NO. IN│ │ │ │ │ - NO.│TABLE │ CU. │ AS. │ SB. │ FE. │ZN. - │ II. │PER CENT.│PER CENT.│PER CENT.│PER CENT.│ - ───┼──────┼─────────┼─────────┼─────────┼─────────┼───── - 1 │ 2 │ 0.0005 │ 0.0003 │ 0.0016 │ 0.0016 │none - 2 │ 3 │ 0.0010 │ 0.0008 │ 0.0107 │ 0.0011 │ “ - ───┴──────┴─────────┴─────────┴─────────┴─────────┴───── - - -TABLE II.—ANALYSES OF BULLION - - ───┬─────────┬─────────┬─────────┬─────────┬──────── - NO.│ FE. │ CU. │ SB. │ SN. │ AS. - │PER CENT.│PER CENT.│PER CENT.│PER CENT.│PER CENT. - ───┼─────────┼─────────┼─────────┼─────────┼───────── - 1 │ 0.0075 │ 0.1700 │ 0.5400 │ 0.0118 │ 0.1460 - 2 │ 0.0115 │ 0.1500 │ 0.6100 │ 0.0158 │ 0.0960 - 3 │ 0.0070 │ 0.1600 │ 0.4000 │ 0.0474 │ 0.1330 - 4 │ 0.0165 │ 0.1400 │ 0.7000 │ 0.0236 │ 0.3120 - 5 │ 0.0120 │ 0.1400 │ 0.8700 │ 0.0432 │ 0.2260 - 6 │ 0.0055 │ 0.1300 │ 0.7300 │ 0.0316 │ 0.1030 - 7 │ 0.0380 │ 0.3600 │ 0.4030 │ │ tr. - ───┴─────────┴─────────┴─────────┴─────────┴───────── - - ───┬─────────┬─────────┬─────────┬─────────┬───────── - NO.│ AG. │ AU. │ PB. │ AG. │ AU. - │PER CENT.│PER CENT.│PER CENT.│OZ. P. T.│OZ. P. T. - ───┼─────────┼─────────┼─────────┼─────────┼───────── - 1 │ 1.0962 │ 0.0085 │ 98.0200 │ 319.7 │ 2.49 - 2 │ 1.2014 │ 0.0086 │ 97.9068 │ 350.4 │ 2.52 - 3 │ 1.0738 │ 0.0123 │ 98.1665 │ 313.2 │ 3.6 - 4 │ 0.8914 │ 0.0151 │ 97.9014 │ 260.0 │ 4.42 - 5 │ 0.6082 │ 0.0124 │ 98.0882 │ 177.4 │ 3.63 - 6 │ 0.6600 │ 0.0106 │ 98.2693 │ 192.5 │ 3.10 - 7 │ 0.7230 │ 0.0180 │ 98.4580 │ 210.9 │ 5.25 - ───┴─────────┴─────────┴─────────┴──────────────────── - - -TABLE III.—ANALYSES OF SLIMES - - ─────────┬─────────┬─────────┬─────────┬─────────┬─────┬────┬───── - FE. │ CU. │ SB. │ SN. │ AS. │ PB. │ZN. │BI. - PER CENT.│PER CENT.│PER CENT.│PER CENT.│PER CENT.│ │ │ - ─────────┼─────────┼─────────┼─────────┼─────────┼─────┼────┼───── - 1.27 │ 8.83 │ 27.10 │ 12.42 │ 28.15 │17.05│none│none - 1.12 │ 22.36 │ 21.16 │ 5.40 │ 23.05 │10.62│ “ │ “ - ─────────┴─────────┴─────────┴─────────┴─────────┴─────┴────┴───── - - -TABLE IV.—ANALYSES OF REFINED LEAD - - ───┬───────┬───────┬───────┬───────┬──────┬───────┬──────┬──────┬───── - │ │ │ │ │ │ │ │ NI, │ - │ CU. │ AS. │ SB. │ FE. │ ZN. │ SN. │ AG. │CO, CD│ BI. - NO.│ PER │ PER │ PER │ PER │ PER │ PER │ OZ. │ PER │ PER - │ CENT. │ CENT. │ CENT. │ CENT. │ CENT.│ CENT. │ P. T.│ CENT.│CENT. - ───┼───────┼───────┼───────┼───────┼──────┼───────┼──────┼──────┼───── - 1 │0.0006 │0.0008 │0.0005 │ │ │ │ │ │ - 2 │0.0003 │0.0002 │0.0010 │0.0010 │ none │ │ │ │ - 3 │0.0009 │0.0001 │0.0009 │0.0008 │ ” │ │ 0.24 │ │ - 4 │0.0016 │ │0.0017 │0.0014 │ │ │ 0.47 │ none │ - 5 │0.0003 │ │0.0060 │0.0003 │ │ │ 0.22 │ │ - 6 │0.0020 │ │0.0010 │0.0046 │ │ │ 0.22 │ none │ - 7 │0.0004 │ none │0.0066 │0.0013 │ none │0.0035 │ 0.14 │ │ - 8 │0.0004 │ │0.0038 │0.0004 │ ” │0.0035 │ 0.25 │ │ - 9 │0.0005 │ │0.0052 │0.0004 │ ” │0.0039 │ 0.28 │ │ - 10 │0.0003 │ none │0.0060 │0.0003 │ ” │0.0049 │ 0.43 │ │ - 11 │0.0003 │ ” │0.0042 │0.0013 │ ” │0.0059 │ 0.32 │ │ - 12 │0.0005 │ ” │0.0055 │0.0009 │ ” │0.0049 │ 0.22 │ │ - 13 │0.0005 │ ” │0.0055 │0.0007 │ ” │0.0091 │ 0.11 │ │ - 14 │0.0004 │ ” │0.0063 │0.0005 │ ” │0.0012 │ 0.14 │ │ - 15 │0.0003 │ ” │0.0072 │0.0003 │ ” │0.0024 │ 0.24 │ │ - 16 │0.0006 │ ” │0.0062 │0.0012 │ ” │0.0083 │ 0.22 │ │ - 17 │0.0006 │ ” │0.0072 │0.0011 │ │0.0080 │ 0.23 │ │ - 18 │0.0006 │ ” │0.0057 │0.0010 │ │0.0053 │ 0.34 │ │ - 19 │0.0005 │ ” │0.0066 │0.0016 │ │0.0140 │ 0.38 │ │ - 19 │0.0005 │ ” │0.0044 │0.0011 │ │0.0108 │ 0.35 │ │ - 20 │0.0004 │ ” │0.0047 │0.0015 │ │0.0072 │ 0.22 │ │ - 20 │0.0004 │ ” │0.0034 │0.0016 │ │ trace │ 0.23 │ │ - 21 │0.0022 │ ” │0.0010 │0.0046 │ none │0.0081 │ 0.38 │ none │ none - ───┴───────┴───────┴───────┴───────┴──────┴───────┴─────────────────── - - -TABLE V.—ANALYSES OF BULLION AND REFINED LEAD - - ──────────────┬───────────┬───────────┬───────────┬────────── - │ AG. │ CU. │ SB. │ PB. - │ PER CENT. │ PER CENT. │ PER CENT. │ PER CENT. - ──────────────┼───────────┼───────────┼───────────┼─────────── - Bullion │ 0.50 │ 0.31 │ 0.43 │ 98.76 - Refined lead │ 0.0003 │ 0.0007 │ 0.0019 │ 99.9971 - ──────────────┴───────────┴───────────┴───────────┴─────────── - - -TABLE VI.—ANALYSES OF BULLION AND REFINED LEAD - - ────────┬──────┬──────┬──────┬──────┬──────┬──────┬─────┬──────┬────── - │ CU. │ BI. │ AS. │ SB. │ AG. │ AG. │ AU. │ FE. │ ZN. - │ PER │ PER │ PER │ PER │ PER │ PER │ PER │ PER │ PER - │ CT. │ CT. │ CT. │ CT. │ CT. │ CT. │ CT. │ CT. │ CT. - ────────┼──────┼──────┼──────┼──────┼──────┼──────┼─────┼──────┼────── - Bullion │0.75 │1.22 │0.936 │0.6832│358.89│ │1.71 │ │ - Refined │ │ │ │ │ │ │ │ │ - lead │0.0027│0.0037│0.0025│0.0000│ │0.0010│none │0.0022│0.0018 - ────────┴──────┴──────┴──────┴──────┴──────┴──────┴─────┴──────┴────── - - -TABLE VII.—ANALYSES OF BULLION, REFINED LEAD AND SLIMES - - ────────────┬─────┬──────┬───────┬───────┬───────┬───────┬──────┬───── - │ │ │ │ │ │ │FE,ZN,│ - │ PB. │ CU. │ AS. │ SB. │ AG. │ AG. │NI,CO.│ BI. - │ PER │ PER │ PER │ PER │ OZ. │ PER │ PER │ - │CENT.│ CENT.│ CENT. │ CENT. │ Per T.│ CENT. │ CENT.│ - ────────────┼─────┼──────┼───────┼───────┼───────┼───────┼──────┼───── - │ │ │ │ │about │ │ │ - Bullion │96.73│0.096 │0.85 │ 1.42 │275[51]│ │ │ - Refined lead│ │0.0013│0.00506│ 0.0028│ │0.00068│0.0027│trace - Slimes (dry │ │ │ │ │ │ │ │ - sample) │ 9.05│1.9 │9.14 │29.51 │9366.9 │ │0.49 │trace - ────────────┴─────┴──────┴───────┴───────┴───────┴───────┴──────┴───── - - -TABLE VIII.—ANALYSES OF BULLION, REFINED LEAD AND SLIMES - - ────────┬─────────┬─────────┬─────────┬─────────┬─────────┬──────── - │ PB. │ CU. │ BI. │ AG. │ SB. │ AS. - │PER CENT.│PER CENT.│PER CENT.│PER CENT.│PER CENT.│PER CENT. - ────────┼─────────┼─────────┼─────────┼─────────┼─────────┼───────── - Bullion │ 87.14 │ 1.40 │ 0.14 │ 0.64 │ 4.0 │ 7.4 - Lead │ │ 0.0010 │ 0.0022 │ │ 0.0017 │ trace - Slimes │ 10.3 │ 9.3 │ 0.52 │ 4.7 │ 25.32 │ 44.58 - ────────┴─────────┴─────────┴─────────┴─────────┴─────────┴───────── - - - - - PART X - - SMELTING WORKS AND REFINERIES - - - - - THE NEW SMELTER AT EL PASO, TEXAS - - (April 19, 1902) - - -In July, 1901, the El Paso, Texas, plant of the Consolidated Kansas -City Smelting and Refining Company[52] was almost completely destroyed -by fire. The power plant, blast-furnace building and blast furnaces -were entirely destroyed, and portions of the other buildings were badly -damaged. The flames were hardly extinguished before steps were taken to -construct a new, modern and enlarged plant on the ruins of the old one, -and on April 15, 1902, nine months after the destruction of the former -plant, the new furnaces were blown in. In rebuilding it was decided to -locate the new power-house at some distance from the other buildings. -The furnaces have all been enlarged, each of the new lead furnaces (of -which there are seven) having about 200 tons daily capacity. These and -the three large copper furnaces have been located in a new position -in order to secure a larger building territory. The entire plant is -modern and up to date in every particular. One of the interesting -features is the substitution of crude oil as fuel in the boiler and -roasting departments. It is intended to use Beaumont petroleum for -the generation of power and the roasting of the ores instead of wood, -coal or coke, and it is expected that a considerable economy will be -effected by this means. - -_Power Plant._—The power plant is complete in all respects. It is a -duplicate plant in every sense of the word, so that it will never be -necessary to shut the works down on account of the failure of any one -piece of machinery. There are seven boilers, having a total of 1250 -h.p. The four blowers are unusually large, having a capacity of 30,000 -cu. ft. of free air per minute. They are direct-connected to three -tandem compound condensing Corliss engines. No belts are used in this -plant, except for driving a small blower of 10,000 cu. ft. capacity, -which will act as a regulator. A large central electric plant has been -installed in the power-house, consisting of two direct-connected, -direct-current generators, mounted on the shafts of two cross-compound -condensing Nordberg-Corliss engines. The current from these generators -is transmitted through the plant, operating sampling works, briquetting -machinery, pumps, hoists, motors, cars, etc., displacing all the -small steam engines and steam pumps used in the old plant. The power -plant is provided with two systems for condensing; one being a large -Wheeler surface condenser, the other a Worthington central-elevated jet -condenser, the idea being to use the surface condenser during a short -period of the year when the water is so bad that it cannot be used in -the boilers. During the remainder of the year the jet condenser is in -service and the surface condenser can be cleaned. The condensed steam -from the surface condenser, with the necessary additional water, goes -back directly to the boilers when the surface condenser is in use. The -power-house is absolutely fireproof throughout, being of steel and -brick with iron and cement floors. It is provided with a traveling -crane, and no expense has been spared to make this, as all other -parts of the plant, complete in every respect. The main conductors -from the generators pass out through a tunnel into a brick and steel -lightning-arrester house, from which point the various distributing -lines go to different parts of the plant. - -_Blast Furnaces._—There are seven large lead furnaces, each having a -capacity of 200 to 250 tons of charge per day, and three large copper -furnaces, each having a capacity of 250 to 300 tons per day. All of -the furnaces are enclosed in one steel fireproof building, the lead -furnaces being at one end and the copper furnaces at the other. Each -of the furnaces has its independent flue system and stack. An entirely -new system of feeding these furnaces has been devised, consisting of -a 6 ton charge car operated by means of a street railroad motor and -controller with third-rail system. The charge cars collect their charge -at the ore beds, lime-rock and coke storage, and are run on to 15 ton -hydraulic elevators. They are then elevated 38 ft. to the top of the -furnaces, traveling over them to the charging doors, through which the -loads are dumped directly into the furnaces. This system permits of two -men handling about 1000 tons per day. The same system and cars are used -for charging the copper furnaces, except that, as these furnaces are -much lower than the lead furnaces, the charge is dropped into a large -hopper, from which it is fed to the copper furnaces by a man on the -copper furnace feed-floor level. - - - - -NEW PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY AT MURRAY, UTAH - - BY WALTER RENTON INGALLS - - (June 28, 1902) - - -Murray is a few miles south of Salt Lake City, with which it is -connected by a trolley line. The new works are situated within a few -hundred yards of the terminus of the latter and in close juxtaposition -to the old Germania plant, which is the only one of the Salt Lake -lead-smelting works in operation at present. The new plant is of -special interest inasmuch as it is the latest construction for -silver-lead smelting in the United States, and may be considered as -embodying the best experience in that industry, the designers having -had access to the results attained at almost all of the previous -installations. It will be perceived, however, that there has been no -radical departure in the methods, and the novelties are rather in -details than in the general scheme. - -The new works are built on level ground; there has been no attempt to -seek or utilize a sloping or a terraced surface, save immediately in -front of the blast furnaces, where there is a drop of several feet -from the furnace-house floor to the slag-yard level, affording room -for the large matte settling-boxes to stand under the slag spouts. -A lower terrace beyond the slag yard furnishes convenient dumping -ground. Otherwise the elevations required in the works are secured by -mechanical lifts, the ore, fluxes and coal being brought in almost -entirely by means of inclines and trestles. - -The plant consists essentially of two parts, the roasting department -and the smelting department. The former comprises a crushing mill -and two furnace-houses, one equipped with Brückner furnaces and the -other with hand-raked reverberatories. The reverberatories are of -the standard design, but are noteworthy for the excellence of their -construction. Similar praise may be, indeed, extended to almost all -the other parts of the works, in which obviously no expense has been -spared on false grounds of economy. The roasting furnaces stand in a -long steel house; they are set at right angles to the longer axis of -the building, in the usual manner. At their feed end they communicate -with a large dust-settling flue, which leads to the main chimney of -the works. The ore is brought in on a tramway over the furnaces and is -charged into the furnaces through hoppers. The furnaces have roasting -hearths only. The fire-boxes are arranged with step-grates and closed -ash-pits, being fed through hoppers at the end of the furnace. The -coal is dumped close at hand from the railway cars, which are switched -in on a trestle parallel with the side of the building, which side is -not closed in. This, together with a large opening in the roof for -the whole length of the building, affords good light and ventilation. -The floor of the house is concrete. The roasted ore is dropped into -cars, which run on a sunken tramway passing under the furnaces. At the -end of this tramway there is an incline up which the cars are drawn -and afterward dumped into brick bins. From the latter it is spouted -into standard-gage railway cars, by which it is taken to the smelting -department. The roasted ore from the Brückner furnaces is handled in a -similar manner. The delivery of the coal and ore to the Brückners and -the general installation of the latter are analogous to the methods -employed in connection with the reverberatories. - -The central feature of the smelting department is the blast-furnace -house, which comprises eight furnaces, each 48 by 160 in. at the -tuyeres. In their general design they are similar to those at the -Arkansas Valley works at Leadville. There are 10 tuyeres per side, a -tuyere passing through the middle of each jacket, the latter being -of cast iron and 16 in. in width; their length is 6 ft., which is -rather extraordinary. The furnaces are very high and are arranged for -mechanical charging, a rectangular brick down-take leading to the dust -chamber, which extends behind the furnace-house. The furnace-house is -erected entirely of steel, the upper floor being iron plates laid on -steel I-beams, while the upper terrace of the lower floor is also laid -with iron plates. As previously remarked, the lower floor drops down a -step in front of the furnaces, but there is an extension on each side -of every furnace, which affords the necessary access to the tap-hole. -The hight of the latter above the lower terrace leaves room for the -large matte settling-boxes, and the matte tapped from the latter runs -into pots on the ground level, dispensing with the inconvenient pits -that are to be seen at some of the older works. The construction of -the blast furnaces, which were built by the Denver Engineering Works -Company, is admirable in all respects. The eight furnaces stand in a -row, about 30 ft. apart, center to center. The main air and water pipes -are strung along behind the furnaces. The slag from the matte-settling -boxes overflows into single-bowl Nesmith pots, which are to be handled -by means of small locomotives. The foul slag is returned by means of a -continuous pan-conveyor to a brick-lined, cylindrical steel tank behind -the furnace-house, whence it is drawn off through chutes, as required -for recharging. - -The charges are made up on the ground level, immediately behind the -furnace-house. The ore and flux are brought in on trestles, whence the -ore is unloaded into beds and the flux into elevated bins. These are -all in the open, there being only two small sheds where the charges are -made up and dumped into the cars which go to the furnaces. There are -two inclines to the latter. At the top of the inclines the cars are -landed on a transferring carriage by which they can be moved to any -furnace of the series. - -The dust-flue extending behind the furnace-house is arranged to -discharge into cars on a tramway in the cut below the ground level. -This flue, which is of brick, connects with the main flues leading to -the chimney. The main flues are built of concrete, laid on a steel -frame in the usual manner, and are very large. For a certain distance -they are installed in triplicate; then they make a turn approximately -at right angles and two flues continue to the chimney. At the proper -points there are large dampers of steel plate, pivoted vertically, for -the purpose of cutting out such section of flue as it may be desired to -clean. Each flue has openings, ordinarily closed by steel doors, which -give access to the interior. The flues are simple tunnels, without -drift-walls or any other interruption than the arched passages which -extend transversely through them at certain places. The chimney is of -brick, circular in section, 20 ft. in diameter and 225 ft. high. This -is the only chimney of the works save those of the boiler-house. - -The boiler-house is equipped with eight internally fired corrugated -fire-box boilers. They are arranged in two rows, face to face. -Between the rows there is an overhead coal bin, from which the coal is -drawn directly to the hoppers of the American stokers, with which the -boilers are provided. Adjoining the boiler-house is the engine-house; -the latter is a brick building, very commodious, light and airy. It -contains two cross-compound, horizontal Allis-Chalmers (Dickson) -blowing engines for the blast furnaces, and two direct-connected -electrical generating sets for the development of the power required -in various parts of the works. A traveling crane, built by the Whiting -Foundry Equipment Company, spans the engine-house. In close proximity -to the engine-house there is a well-equipped machine shop. Other -important buildings are the sampling mill and the flue-dust briquetting -mill. - -A noteworthy feature of the new plant is the concrete paving, laid on a -bed of broken slag, which is used liberally about the ore-yard and in -other places where tramming is to be done. The roasting-furnace houses -are floored with the same material, which not only gives an admirably -smooth surface, but also is durable. The whole plant is well laid out -with service tramways and standard-gage spur tracks; the intention has -been, obviously, to save manual labor as much as possible. - - - - - THE MURRAY SMELTER, UTAH[53] - - BY O. PUFAHL - - (May 26, 1906) - - -This plant has been in operation since June, 1902. It gives employment -to 800 men. The monthly production consists of about 4000 tons of -work-lead and 700 tons of lead-copper matte (12 per cent. lead, 45 per -cent. copper). The work-lead is sent to the refinery at Omaha; the -matte to Pueblo, Colo. Most of the ores come from Utah; but in addition -some richer lead ores are obtained from Idaho, and some gold-bearing -ores from Nevada. - -For sampling the Vezin apparatus is used, cutting out one-fifth in -each of three passes, crushing intervening, the sample from the third -machine being 1/625 of the original ore; after further comminution of -sample in a coffee-mill grinder, it is cut down further by hand, using -a riffle. The final sample is bucked down to pass an 80-mesh sieve, but -gold ores are put through a 120-mesh. - -The steps in the smelting process are as follows: Roasting the poorer -ores in reverberatory furnaces and in Brückner cylinders. Smelting -raw and roasted ores, mixed, in water-jacketed blast furnaces, -for work-lead and lead-copper matte, the latter containing 15 per -cent. lead and 10 to 12 per cent. copper. Roasting the ground -matte, containing 22 per cent. of sulphur, down to ¾ per cent. in -reverberatory furnaces. Smelting the roasted matte together with acid -flux in the blast furnace for a matte with 45 per cent. copper and 12 -per cent. lead. - -Only the pyritic ores are roasted in Brückner furnaces, the lead ores -and matte being roasted in reverberatory furnaces. Each of the 20 -Brückner furnaces, which constitute one battery, roasts 8 to 12 tons -of ore in 24 hours down to 5½ to 6 per cent. sulphur, with a coal -consumption of two tons. The charge weighs 24 tons. The furnaces make -one turn in 40 minutes. To increase the draft and the output, steam -at 40 lb. pressure is blown in through a pipe; this has, however, -resulted in increasing the quantity of flue dust to 10 to 15 per cent. -of the ore charged. Ten furnaces are attended by one workman with one -assistant, working in eight-hour shifts. For firing and withdrawing the -charge five men are required. - -The gases from the Brückners and reverberatory furnaces pass into a -dust-flue 14 × 14 ft. in section and 600 ft. long, built of brickwork, -with concrete vault; in the stack (225 ft. high, 20 ft. diameter) they -unite with the shaft-furnace gases, the temperature of which is only 60 -deg. - -There are 12 reverberatory furnaces with hearths 60 ft. long and 16 -ft. broad. They roast 14 tons of ore (or 13 tons of matte) in 24 hours -down to 3½ to 4 per cent. sulphur, consuming 32 to 34 per cent. of coal -figured on the weight of the charge. There are 12 working doors on each -side. The small coal (from Rock Springs, Wyoming), which is burnt on -flat grates, contains 5 per cent. ash and 3 to 5 per cent. moisture. -The roasted product is dumped through an opening in the hearth, -ordinarily kept closed with an iron plate, into cars which are raised -by electricity on a self-acting inclined plane. Their content is then -tipped over into a chute and cooled by sprinkling with water. From here -the roasted matte is conveyed to the blast furnace in 30-ton cars. The -roasted ore is tipped into the ore-bins. - -There are eight blast furnaces, 48 × 160 in. at the tuyeres, of which -there are 10 on each of the long sides. The hight from the tuyeres to -the gas outlet is 20 ft., thence to the throat 6 ft.; the distance -of the tuyeres from the floor is 4 ft. The base is water-cooled. The -water-jackets of the furnace are 6 ft. high. The tuyeres (4 in.) -are provided with the Eilers automatic arrangement for preventing -the furnace gases entering the blast pipes. The blast pressure is -34 oz. The furnaces are furnished with the Arents lead wells; the -crucible holds about 30 tons of lead. The slag and the matte run into -a brick-lined forehearth (8 × 3 ft., 4 ft. deep), from which the slag -flows into pots holding 30 cu. ft., while the matte is tapped off into -flat round pans mounted on wheels. - -The charge is conveyed to the feed-floor by electricity. The furnace -charge is 8000 lb. and 12 per cent. coke, with 30 per cent, (figured on -the weight of the charge) of “shells” (slag). Occasionally as much as -230 tons of the (moist) charge, exclusive of coke and slag, has been -handled by one furnace in 24 hours. During one month (September, 1904) -40,000 tons of charge were worked up, corresponding to a daily average -of 166 tons per furnace. - -The lead in the charge runs from 13 to 14 per cent. on an average. The -limestone, which is added as flux, is quarried not far from the works. -The coke used is in part a very ordinary quality from Utah; in part a -better quality from the East, with 9 to 10 per cent. ash. The matte -amounts to 10 per cent. The slag contains 0.6 to 0.7 per cent. lead and -0.1 to 0.15 per cent. copper. The slag has approximately the following -composition: 36 per cent. silica, 23 per cent. iron (corresponding to -29.57 per cent. FeO), 23 per cent. lime, 3.8 per cent. zinc and 4 per -cent. alumina. - -The work-lead is transferred while liquid from the furnaces to kettles -of 30 tons capacity, in which it is skimmed, and thence cast in molds -through a Steitz siphon. First, however, a 5.5 lb. sample is taken -out by means of a special ladle, and is cast into a plate. From this -samples of 0.5 a.t. are punched out at four points for the assay of the -precious metals. - -For the purpose of precipitating the flue dust, the blast-furnace gases -are passed into brickwork chambers in which a plentiful deposition of -the heavier particles takes place. From here the gases go through an L -pipe of sheet iron, 18 ft. in diameter, to the Monier flues, which have -a cross-section of 256 sq. ft. and a total length of 2000 ft. A small -part of the flues is also built of brick. The gases unite with the hot -roaster gases just before entering the 225 ft. chimney. In the portion -of the blast-furnace dust first precipitated the silver runs 22 oz. per -ton, while that recovered nearer the stack contains only 8 oz. The flue -dust is briquetted with a small proportion of lime, and, after drying, -is returned to the blast furnaces. - - - - - THE PUEBLO LEAD SMELTERS[54] - - BY O. PUFAHL - - (May 12, 1906) - - -At the Pueblo plant, ores containing over 10 per cent. lead are not -roasted, but are added raw to the charge. For such material as requires -roasting there are in use five Brückner furnaces. The charge is 24 tons -for 48 to 60 hours; the furnaces make one revolution per minute and -roast the ore down to 6 per cent. sulphur. There are also two O’Harra -furnaces, each roasting 25 tons daily, and 10 reverberatory furnaces 75 -ft. in length, each roasting 15 tons of ore daily down to 4 per cent. -sulphur. - -The charge for smelting is prepared from roasted ore, together with -Idaho lead ore, Cripple Creek gold ore, briquetted flue dust, slag -and limestone. There are seven water-jacketed furnaces, which smelt, -each, 150 tons of charge per day. The furnaces have 18 tuyeres, blast -pressure 34 oz., cross-section at the tuyeres 48 × 148 in. They are -charged mechanically by a car of 4 tons’ capacity. - -The output of lead is 11 to 15 tons per furnace. The matte, which -is produced in small quantity, contains 8 to 12 per cent. lead and -the same percentage of copper. It is crushed by rolls, roasted in -reverberatory furnaces, and smelted with ores rich in silica. The matte -resulting at this stage, running 45 to 50 per cent. in copper, is -shipped to be further worked up for blister copper. - -The work-lead is purified by remelting in iron kettles, the cupriferous -dross being pressed dry in a Howard press, and sent to the blast -furnaces. The work-lead is sent to the refineries at Omaha, Neb., or -Perth Amboy, N. J. - -To collect the flue dust the waste gases are passed through long brick -flues. The chimneys are 150 to 200 ft. high, and 15 ft. in diameter. -They stand 75 ft. above the ground level of the blast furnaces. The -comparatively small proportion of flue dust produced (0.9 per cent. of -the charge) is briquetted, together with fine ore and 5 per cent. of -a thick paste of lime. For this purpose a White press is used, which -makes six briquets at a time, and handles 10 tons per hour. - -According to a tabulation of the results of five months’ running, the -proportion of flue dust at several works of the American Smelting and -Refining Company is as follows: - - Globe Plant, Denver 0.5% of the charge. - Pueblo Plant, Pueblo 0.9% ” ” ” - Eilers’ Plant, Pueblo 0.5% ” ” ” - East Helena Plant, Helena 0.3% ” ” ” - Arkansas Valley Plant, Leadville 0.2% ” ” ” - Murray Plant, Murray, Utah 1.2% ” ” ” - -The fuel used is of very moderate quality. The coke (from beehive -ovens) carries up to 17 per cent. ash, the coal 10 to 18 per cent. The -monthly production is 2300 tons of work-lead and 150 tons of copper -matte (45 to 50 per cent. copper). - -At the Eilers plant all sulphide ores, except the rich Idaho ore, are -roasted down to 5 to 7 per cent. S in 15 reverberatory furnaces, 60 to -70 ft. in length, each furnace roasting 15 tons per 24 hours, in six -charges. - -The flue dust is briquetted together with fine Cripple Creek ore, -pyrites cinder from Argentine, Kan., Creede ores rich in silica and -10 per cent. lime. The residue from the zinc smeltery (U. S. Zinc -Company), which is brought to this plant (600 tons a month containing -nearly 10 per cent. lead), is taken direct to the blast furnaces. -Of the latter there are six, each with 18 tuyeres, which handle per -24 hours 160 to 180 tons of charge, containing on an average 10 per -cent. of lead in the ore, with 10 per cent. of coke, figured on the -charge. The average monthly production of a furnace is about 360 tons -of work-lead, which is purified at the Pueblo plant. The furnaces -are charged by hand. Of the slag, 30 per cent., as shells, etc., is -returned to the charge. The monthly production of work-lead is 2000 -tons, carrying 150 oz. of silver and 2 to 6 oz. of gold per ton. - -The matte amounts to about 8.3 per cent., and contains 12 per cent. -copper. It is concentrated up to 45 per cent. Cu, which is shipped (150 -tons a month) for smelting to blister copper. - - - - -THE PERTH AMBOY PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY[55] - - BY O. PUFAHL - - (January 27, 1906) - - -These works were erected in 1895 by the Guggenheim Smelting Company. -They are situated on Raritan Bay, opposite the southern point of Staten -Island, in a position offering excellent facilities for transportation -by land and by water. The materials worked up are base lead bullion -and crude copper, containing silver and gold, chiefly drawn from the -company’s smelteries in the United States and Mexico. Silver ore is -received from South America. The ores and base metals from Mexico and -South America are brought to Perth Amboy by the company’s steamships -(American Smelters Steamship Company). - -_Ore Smelting._—The silver ore from South America (containing antimony -and much silver, together with galena, iron and copper pyrites) is -crushed by rolls and is roasted down from 26 per cent. to 3 per cent. -S in 11 reverberatory furnaces, 70 ft. long and 15 ft. wide (inside -dimensions). It is then mixed with rich galena from Idaho, pyrites -cinder, litharge, copper skimmings, and residues from the desilverizing -process, together with limestone, and is smelted for work-lead and -lead-copper matte in three water-jacketed furnaces, using 12 per cent. -coke, figured on the ore in the charge. Of these furnaces one has 12 -tuyeres; it measures 42 × 96 in. in cross-section at the tuyeres, and -6 ft. 3 in. by 8 ft. at the charging level. The hight of charge is 16 -ft. The other two furnaces have 16 tuyeres each, their cross-section at -the tuyeres being 44 in. by 128 in., at the charging level 6 ft. 6 in. -by 12 ft., and hight of charge 16 ft. The furnaces are operated at a -blast pressure of 35 oz. per square inch. The temperature of the gases -at the throat is 140 deg. F. (60 deg. C.) measured with a Columbia -recording thermometer, which works very well. These furnaces reduce, -respectively, 100 to 120 and 130 to 140 tons of charge per 24 hours; -they are also used for concentrating roasted matte. - -_Copper Refining._—The crude copper is melted in two furnaces of 125 -tons aggregate daily capacity, and is molded into anodes by Walker -casting machines. Twenty-six anodes are lifted out of the cooling -vessel at a time, and are taken to the electrolytic plant. - -The electrolytic plant comprises two systems, each of 408 vats. The -current is furnished by two dynamos, each giving 4700 amperes at 105 -volts. The cathodes remain in the bath for 14 days. The weight of the -residual anodes is 15 per cent. - -The anode mud is swilled down into reservoirs in the cellar as at -Chrome (De Lamar Copper Refining Company), is cleaned, dried and -refined in a similar manner. - -For melting the cathodes there are two reverberatory furnaces of -capacity for 75 tons per 24 hours. The wire-bars and ingots are cast -with a Walker machine. About 3200 tons of refined copper are produced -per month. - -_Copper Sulphate Manufacture._—The lyes withdrawn from the electrolytic -process are worked up into copper sulphate, shot copper being added. -This latter is prepared in a reverberatory furnace from matte obtained -as a by-product in working up the lead. About 200 tons of copper -sulphate are thus produced per month; the process used is the same as -at the Oker works. Lower Harz, Germany. The crystals are rinsed, dried -and packed in strong wooden barrels. - -_Lead Refining._—The working up of the Mexican raw lead is carried -out under the supervision of the customs officers. The lead, which is -imported duty free, must be exported again. From each bar a sample is -cut from above and below by means of a punch entering half way into the -bar. For refining the lead there are four reverberatory furnaces of 60 -tons capacity, with hearths 17 ft. 9 in. by 12 ft. 6 in., a mean depth -of 14 in., and a grate area of 2 ft. 6 in. by 6 ft.; in addition to -these there is a furnace of 80 tons capacity with a hearth 19 ft. 7½ -in. by 9 ft. 6 in., a mean depth of 18 in., and grate area of 3 ft. by -6 ft. - -For desilverizing the softened lead there are five kettles, each of -60 tons capacity, 10 ft. 3 in. diameter and 39 in. depth. The zinc -is stirred in with a Howard mechanical stirrer and the zinc scum is -pressed dry in a Howard press, which gives a very dry scum. The latter -is then, while still warm, readily hammered into pieces for the retorts. - -The desilverized lead is refined in five reverberatory furnaces, of -which four take a charge of 50 tons each, and one of 65 tons. The -production of desilverized lead is 5000 to 5500 tons a month. - -The distillation of the zinc crusts is carried out in 18 oil-fired -Faber du Faur tilting furnaces. Each retort receives a charge of -1200 lb. of broken-up crust and a little charcoal. The distillation -lasts 6 to 7 hours. Fifty gallons of petroleum residues are consumed -per charge. The oil is blown into the furnace with a compressed -air atomizer. After withdrawing the condenser, which runs on a -traveling support, the argentiferous lead is poured directly from -the tilted retort into an English cupel furnace. Seven such furnaces -(magnesia-lined, with movable test) are in use, of which each works -up 4.5 to 5 tons of retort metal in 24 hours. The furnaces are -water-jacketed. The blast is introduced by the aid of a jet of steam. -Three tons of coal are used per 24 hours. - -_Gold and Silver Parting._—The doré bars are cast into anodes for -electrolytic parting by the Moebius process. The plant consists of 144 -cells in 24 divisions. The mean composition of the electrolytic bath is -said to be as follows: 10 per cent. free nitric acid, 17 grams silver, -and 35 to 40 grams copper per liter. The current is furnished by a 62 -k.w. dynamo. One cell consumes 260 amp. at 1.75 volts. One k.w. gives -a yield of 1600 oz. fine silver per 24 hours. The daily production -of silver is almost 100,000 oz., and is exceeded at no other works. -About $3,000,000 worth of metal is always on hand in the different -departments. - - - - - THE NATIONAL PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY[56] - - BY O. PUFAHL - - (April 14, 1906) - - -This plant, at South Chicago, Ill., refines base lead bullion. It -comprises four reverberatory furnaces, of which one takes a charge of -100 tons, one 80 tons, and the other two 30 tons each; one of the small -furnaces is being torn down, and a 120 ton furnace is to be built in -its place. The furnaces are fired with coal from Southern Illinois, -which contains 11 per cent. of ash. - -In softening the bullion, the time for each charge is 10 hours. The -first portion tapped consists of dross rich in copper, which is -followed by antimonial skimmings and litharge. - -The copper dross is melted up in a small reverberatory furnace, -together with galena from Wisconsin (containing 80 per cent. lead), -for work-lead and lead-copper matte, the latter containing about 35 -per cent. of copper; this matte is enriched to 55 per cent. copper -by the addition of roasted matte, and is finally worked up for crude -copper (95 per cent.) in a reverberatory furnace. All the copper so -produced is used in the parting process for precipitating the silver. -The antimonial skimmings are smelted in a reverberatory furnace, -together with coke cinder, for lead and a slag rich in antimony, which -is reduced to hard lead (27 per cent. antimony, 0.5 per cent. copper, -0.5 per cent. arsenic) in a small blast furnace, 14 ft. high, which has -8 tuyeres. - -The softened lead is tapped off into cast-iron desilverizing pots, -which usually outlive 200 charges; in isolated cases as many as -300. For desilverizing, zinc from Pueblo, Colo., is added in two -instalments, being mixed in by means of a Howard stirrer. After the -first addition there remains in the lead 7 oz. of silver per ton; -after the second only 0.2 oz. The first scum is pressed in a Howard -press and distilled; the second is ladled off and is added to the next -charge. The Howard stirrer is driven by a small steam engine suspended -over the kettle; the Howard press by compressed air. - -For distilling zinc scum, 12 Faber du Faur tilting retorts, heated with -petroleum residue, are used. The argentiferous lead (with 9.6 per cent. -silver) is transferred from the retort to a pan lined with refractory -brick, which is wheeled to the cupelling hearth and raised by means of -compressed-air cylinders, so as to empty its molten contents through a -short gutter upon the cupelling hearth. The cupelling hearths are of -the water-cooled English type, and are heated by coal with under-grate -blast. The cast-iron test rings, with reinforcing ribs, are made in two -pieces, slightly arched and water-cooled; they are rectangular, with -rounded corners, and are mounted on wheels. The material of the hearth -is marl. - -Argentiferous lead is added as the operation proceeds, and finally the -doré bullion is poured from the tilted test into thick bars (1100 oz.) -for parting. - -The desilverized lead is refined in charges of 28 tons (4 to 5 hours) -and 80 to 90 tons (8 to 10 hours), introducing steam through four to -eight half-inch iron pipes. The first skimmings contain a considerable -proportion of antimony and are therefore added to the charge when -reducing the antimonial slags in the blast furnace. The litharge is -worked up in a reverberatory furnace for lead of second quality. The -refined lead is tapped off into a kettle, from which it is cast into -bars through a siphon. - -The parting of the doré bullion is carried out in tanks of gray cast -iron, in which the solution is effected with sulphuric acid of 60 deg. -B. The acid of 40 deg. B. condensed from the vapors is brought up to -strength in leaden pans. In a second larger tank, which is slightly -warmed, a little gold deposits from the acid solution of sulphates. -The solution is then transferred (by the aid of compressed air) to the -large precipitating tank, and diluted with water. It is here heated -with steam, and the silver is rapidly precipitated by copper plates -(125 plates 18 × 8 × 1 in.) suspended in the solution from iron hooks -covered with hard lead. After the precipitation, the vitriol lye is -siphoned off, the silver is washed in a vat provided with a false -bottom, is removed with a wooden shovel, and is pressed into cakes 10 × -10 × 6 in. - -The refining is finished on a cupelling hearth fired with petroleum -residue, adding saltpeter, and removing the slag by means of powdered -brick. After drawing the last portion of slag the silver (0.999 fine) -is kept fused under a layer of wood-charcoal for 20 minutes, and is -then cast into iron molds, previously blackened with a petroleum flame. -The bars weigh about 1100 oz. - -The gold is boiled with several fresh portions of acid, is washed and -dried, and finally melted up with a little soda in a graphite crucible. -It is 0.995 fine. - -The lye from the silver precipitation, after clearing, is evaporated -down to 40 deg. B. in leaden pans by means of steam coils, and is -transferred to crystallizing vats. The first product is dissolved -in water, the solution is brought up to 40 deg. B. strength, and is -allowed to crystallize. The purer crystals so obtained are crushed, and -are washed and dried in centrifugal apparatus; they are then sifted and -packed in wooden casks in two grades according to the size of grain. -The very fine material goes back into the vats. From the first strongly -acid mother liquor, acid of 60 deg. B. is prepared by concentrating in -leaden pans, and this is used for the parting operation. - - - - -THE EAST HELENA PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY[57] - - BY O. PUFAHL - - (April 28, 1906) - - -The monthly production of these works is about 1500 tons of base -bullion (containing 150 oz. Ag and 4 to 6 oz. Au per ton), and 200 tons -of 45 per cent. copper matte. The base bullion is shipped to South -Chicago, the matte to Pueblo. - -The ore-roasting is done in two batteries of eight reverberatory -furnaces and 16 Brückner furnaces, the resulting product containing on -an average 20 per cent. lead and 3 per cent. sulphur. The charge for -the blast furnaces consists of roasted ore, rich galena, argentiferous -red hematite, briquetted flue dust, slag (shells) from the furnace -itself, lead skimmings, scrap iron and limestone. - -Four tons of the charge are dumped over a roller into a low car, which -is then drawn up an inclined plane to the charging gallery by an -electric motor and is then dumped into the furnace. - -The two rectangular blast furnaces (Eilers’ type) have eight tuyeres on -each of their longer sides and cast-iron water-jackets of 6 ft. hight. -The blast is delivered at a pressure of 40 oz. The lead is drawn off -through a siphon tap into a cooling kettle. The furnace has a large -forehearth for separating the matte and the slag. The slag is received -by a two-pot Nesmith truck, having an aggregate capacity of 14 cu. ft. -These trucks are hauled to the dump by an electric locomotive. The -shells are returned to the furnace with the charge. - -The matte (with about 6 per cent. Cu and the same percentage of lead) -is tapped off into iron molds and after cooling is crushed to 0.25-in. -size, to be roasted in the reverberatory furnaces and smelted up -together with roasted ore for a 15 per cent. matte. The latter is -crushed, roasted and separately smelted together with silicious ore -for 45 per cent. matte, which is then sent to Pueblo to be worked up -into blister copper. The small quantity of speiss which is formed is -broken up and returned to the blast furnaces with the charge. The slag -contains 0.5 to 0.8 per cent. lead and 0.5 oz. silver per ton. - - - - - THE GLOBE PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY[58] - - BY O. PUFAHL - - (May 5, 1905) - - -This plant produces 1800 tons of base bullion per month and 200 tons -of lead-copper matte containing 45 to 52 per cent. of copper. The ores -smelted are mostly from Colorado, but include also galena from the Cœur -d’Alene and other supplies. The limestone is quarried 14 miles from -Denver; coke and coal are brought from Trinidad, Colo. - -All sulphides, except the slimes, concentrates and the rich Idaho ores, -are roasted. For roasting there are: - -(1) Fifteen reverberatory furnaces, five of which measure 60 × 14 ft., -and the other ten 80 × 16 ft. externally. In 24 hours these roast six -charges of 4400 lb. (average) of moist ore (2.15 tons of dry ore) from -28 to 30 per cent. down to 3 to 4 per cent. sulphur. Each furnace is -attended by three men working in 12-hour shifts; the stoker earns -$2.75; the roasters, $2.30. - -(2) Two Brown-O’Harra furnaces, 90 ft. long, with two hearths, and a -small sintering furnace attached. They have three grates on each long -side, and each roasts 26 tons of ore in 24 hours down to ¾ per cent. -sulphur. - -(3) Twelve Brückner furnaces, each taking 24 tons’ charge, with -under-grate blast, the air being fed into the cylinders by a steam jet. -According to the zinc content of the ores the roasting operation lasts -70 to 90 hours, the furnace making one revolution per hour. The roasted -product from the Brückner furnaces is pressed into briquets, together -with fine ore, flue dust and lime. - -The smelting is carried out in seven blast furnaces, with 16 tuyeres, -blast at 2-lb. pressure, hight of furnace 18 ft. 6 in., section at -the tuyeres 42 × 144 in. The charge is 120 to 150 tons exclusive of -slag and coke. The slag and the matte are tapped off together into -double-bowl Nesmith cars, which are hauled, by an electric locomotive, -to a reverberatory furnace (hearth 20 × 12 ft.) in which they are kept -liquid, for several hours, in charges of 14 to 15 tons, in order to -effect complete separation. A little work-lead is obtained in this -operation, while the matte is tapped off into cast-iron pans of one -ton capacity, and the slag, 0.5 to 0.6 per cent. lead, 0.6 to 0.7 oz. -silver, is removed in 5-ton pots to the dump. - -The matte is broken up, crushed to 0.25 in. size, roasted in the -reverberatory furnaces, smelted for a 45 to 52 per cent. copper matte, -which is shipped to be further worked up into blister copper. The crude -matte contains 10 to 12 per cent. copper, 12 to 15 per cent. lead, 40 -oz. silver and 0.05 oz. gold. - -From the siphon taps of the blast furnaces the work-lead is transferred -to a cast-iron kettle of 33 tons’ capacity. Here the copper dross -is removed, the metal is mixed by introducing steam for 10 minutes, -sampled, and the lead is cast into bars through siphons. It contains -about 2 per cent. antimony, 200 oz. silver and 8 oz. gold. This product -is refined at Omaha. - -The blast-furnace gases pass through a flue 1200 ft. long, and enter -the bag-house, in which they are filtered through 4000 cotton bags 30 -ft. long and 18 in. in diameter. These bags are shaken every 6 hours. -The material which falls to the floor is burnt where it lies, sintered -and returned to the blast furnaces. - -In the engine house there are four Connersville blowers, two of which -are No. 8 and two of No. 7 size. Each blast furnace requires 45,000 cu. -ft. of air a minute. - -The works give employment to 450 men, whose wages (for 10-to 12-hour -shifts) are $2 to $3. - - - - - LEAD SMELTING IN SPAIN - - BY HJALMAR ERIKSSON - - (November 14, 1903) - - -A few notes, gathered during a couple of years while I was employed -at one of the large lead works in the southeastern part of Spain, are -of interest, not as showing good work, but for comparing the results -obtained in modern practice with those obtained by what is probably the -most primitive kind of smelting to be found today. The plant about to -be described may serve as a general type for that country. As far as I -know, the exceptions are a large plant at Mazarron, fully up to date -and equipped with the most modern improvements in every line; a smaller -plant at Almeria, also in good shape, and the reverberatory smelting of -the carbonates at Linares. It should be kept in mind, however, that the -conditions are peculiar, iron and machinery being very expensive and -manual labor very cheap. - -[Illustration: FIG. 41.—Spanish Lead Blast Furnace.] - -About 4 ft. above the tuyeres the furnace is built of uncalcined brick -made of a black graphitic clay found in the mines near by; the upper -part is of common red brick. The entire cost of one furnace does not -reach $100. The flue leads to a main gallery 3.5 by 7 ft., which goes -down to the ground, and extends several times around a hill, the -chimney being placed on the top of the hill, considerably above the -furnace level. The gallery is about 10,000 ft. long, and is laid down -in the earth, with the arched roof just emerging. It is all built of -rough stone, the inside being plastered with gypsum. The furnace has -three tuyeres of 3 in. diameter. The blast pressure is generally 4 to -6 in. of water. Neither feeding floor nor elevators are used, only a -couple of scaffolds, the charge being lifted up gradually by hand in -small convenient buckets made of sea-grass. When charging the furnace, -coke is piled up in the center, and the mixture of ore, fluxes and slag -is charged around the walls. The slag and matte are left to run out -together on an inclined sand-bed. The matte, flowing more quickly, goes -further and leaves the slag behind, but the separation thus obtained -is, of course, very unsatisfactory. The charge mixture is weighed and -made for each furnace every morning. When it is all put through, the -furnace is run down very low, without any protecting cover on the top; -several iron bars are driven through the furnace at the slag-tap level, -for holding up the charge; the lead is all tapped out; a big hole is -made in the crucible for the purpose of cleaning it out; all accretions -are loosened with a bar; the hole is closed with mud of the graphitic -clay; bars are removed, when the crucible is filled with coke from the -center and the charging is continued. In this way a furnace can be kept -running for any length of time, but at a great loss of heat, and with a -great increase of flue dust. - -The current practice, in many parts of Spain, is to run the same number -of ore-smelting and of matte-smelting furnaces. All the slag and the -raw matte, produced by the ore-smelting furnaces, is re-smelted in the -matte furnaces, together with some dry silver ores. No lead at all is -produced in the matte furnaces, only a matte containing up to 150 oz. -silver per ton and 25 to 35 per cent. of the lead charged on them. This -rich matte is calcined in kilns, and smelted together with the ore -charge. - -The ores we smelted were galena ranging from 5 to 83 per cent. lead -and about 250 oz. silver per ton of lead; dry silver ores containing -up to 120 oz. silver per ton, and enough of the Linares carbonates for -keeping the silver below 120 oz. per ton in the lead. The gangue of the -galena was mainly iron carbonate. Most of that ore was hand picked and -of nut size. Machine concentrates with more than 30 per cent. lead or -containing much pyrite were calcined; everything else was smelted raw. -The flux exclusively used, before I came, was carbonate of iron, which, -by the way, was considered a “cure-for-all.” The slag analyses showed: - - CaO, below 4 per cent. - FeO, above 45 per cent. - SiO₂, about 30 per cent. - BaO, 5 to 10 per cent. - Al₂O₃, 5 to 10 per cent. - Pb, by fire assay, 0.75 to 2.5 per cent. - Ag, by fire assay, 2 to 3 oz. per ton. - -The specific gravity of the slag was about 5, or practically the same -as that of the matte. The output of metallic lead was about 70 per -cent.; of silver, 84 per cent. The working hight of the furnaces—tuyere -level to top of charge—was at that time only 7 ft., and I was told that -it had been still lower before. - -To the working hight of the furnaces was added 2 ft., simply by putting -up the charging doors that much. A very good limestone was found just -outside the fence around the plant. Enough limestone was substituted -for the iron carbonate, to keep the lime up to 12 per cent. in the -slag, reducing the FeO to below 35 per cent. and the specific gravity -to below four. - -The result of these alterations was an increase in the output of -metallic lead, from 76 to 85 per cent.; of silver from 84 to 90 per -cent.; a comparatively good separation of slag and matte, and a slag -running about 0.5 to 0.75 per cent. Pb and 1.5 oz. Ag per ton. - -Owing to the great extent of the gallery, and the consequent good -condensation of the flue dust, the total loss of lead and silver was -much smaller than would be expected; in no case being found above 4 per -cent. - -The composition of the charge was 55 per cent. ore and roasted matte, -13 per cent. fluxes, and 32 per cent. slag. Coke used was 11 per cent. -on charge, or 20 per cent. on ore smelted. Each furnace put through 10 -to 15 tons of charge, or 7 tons of ore, in 24 hours. Eight men and two -boys were required for each furnace, including slag handling and making -up of the charge. The cost of smelting was 17 pesetas per ton of ore, -which at the usual premium (£1 = 34 pesetas = $4.85) equals $2.43. This -cost is divided as follows: - - Coke $1.47 - Fluxes 0.04 - Labor 0.65 - Coal for power 0.10 - General expenses 0.17 - ————- - Total $2.43 - -This $2.43 per ton includes all expenses of whatever kind. The iron -carbonate flux contained lead and silver, which was not paid for. The -fluxes are credited for the actual value of this lead and silver. -Without making this discount, the cost of flux would amount to 26c. per -ton, making the entire smelting cost come to $2.65. As an explanation -of the low cost of labor, it may be noted that the wages were, for -the furnace-man, 2.25 pesetas, or 32c. a day; for the helpers, 1.75 -pesetas, or 25c. a day. - -The basis for purchasing the galena ore may here be given, reduced to -American money; lead and silver are paid for according to the latest -quotations for refined metals given by the _Revista Minera_, published -at Cartagena. (The quotations are the actual value in Cartagena of the -London quotations.) - -The following discounts are made: 5 per cent. for both silver and -lead; $6.40 per ton on ore containing 7 per cent. Pb and below; this -rises gradually to a discount of $7.75 per ton of ore containing 30 per -cent. Pb and above. - -The transportation is paid by the purchaser and amounts to about $1.20 -per ton of ore. - -The dry silver ores were cheaper than this and the lead carbonates much -more expensive. - - - - - LEAD SMELTING AT MONTEPONI, SARDINIA[59] - - BY ERMINIO FERRARIS - - (October 28, 1905) - - -In dressing mixed lead and zinc carbonate ores by the old method of -gradual crushing with rolls, middling products were obtained, which -could be further separated only with much loss. Inasmuch as the losses -in the metallurgical treatment of such mixed ore were reckoned to be -less than in ore dressing, these between-products at Monteponi were -saved for a number of years, until there should be enough raw material -to warrant the erection of a small lead and zinc smeltery. - -In 1894 the lead smeltery in Monteponi was put in operation; in 1899 -the zinc smeltery was started. At about the same time the reserves of -lead ore were exhausted, and the lead plant then began to treat all the -Monteponi ores and a part of those from neighboring mines. - -As will be seen from the plan (Fig. 42), the smelting works cluster -in terraces around the mine shaft, covering an area of about 3000 sq. -m. (0.75 acre); the ore stocks and the pottery of the zinc works are -located in separate buildings. - -During the first years of working, the slag had purposely been kept -very rich in zinc, in the hope of utilizing it later for the production -of zinc oxide. It had an average zinc content of 16.80 per cent., or 21 -per cent. of zinc oxide, with about 32 per cent. SiO₂, 25 per cent. -FeO, and 14 per cent. lime. According to the recent experiments, this -slag can very well be used for oxide manufacture, in connection with -calamine rich in iron. The slag made at the present time has only 15 -per cent. ZnO; 25 per cent. SiO₂; 16 per cent. CaO; 3 per cent. MgO; -33 per cent. FeO; 2.5 per cent. Al₂O₃, and 2 per cent. BaO, and -small quantities of alkalies, sulphur and lead (1 to 1.5 per cent). - -The following classes of ore are produced at Monteponi: - -1. Lead carbonates, with a little zinc oxide; these ores are screened -down to 10 mm. The portion held back by the screen is sent straight -to the shaft furnaces; the portion passing through is either roasted -together with lead sulphides, or is sintered by itself, according to -circumstances. - -2. Dry lead ores, mostly quartz, with 10 to 15 per cent. lead, which -are mixed for smelting with the lead carbonates. - -[Illustration: FIG. 42.—General Plan of Works.] - -3. Lead sulphides, which are crushed fine and roasted dead. Quartz -sand is added in the roasting, in order to decompose the lead sulphate -and produce a readily fusible silicate; as quartz flux, fine sand from -the dunes on the coast is used. This is a product of decomposition of -trachyte, and contains 88 per cent. of silica, together with alkalies -and alumina. The roast is effected in two hand-raked reverberatory -furnaces, 18 m. long, which turn out 12,000 kg. of roasted ore in -24 hours, consuming 1800 kg. of English cannel coal, or 2400 kg. -of Sardinian lignite. There is also a third reverberatory furnace, -provided with a fusion chamber, which is used for roasting matte and -for liquating various secondary products. - -The charge for the shaft furnace, as a rule, consists of 50 per cent. -ore (crude and roasted), 20 per cent. fluxes and 30 per cent. slag -of suitable origin. The fluxes used are limestone from the mine, -containing 98 per cent. CaCO₃, and limonite from the calamine -deposits. This iron ore contains 48 per cent. Fe, not more than 4 per -cent. Zn, a little lead and traces of copper and silver. - -A shaft furnace will work up a charge of 60 tons, equal to 30 tons of -ore, in 24 hours, with a coke consumption of 12 per cent. of the weight -of the charge and a blast pressure of 50 mm. of mercury. There are -three furnaces, of which two are used alternately for smelting lead -ores, while one smaller furnace serves for smelting down products, such -as hard lead, copper matte and copper bottoms. - -[Illustration: FIG. 43.—Elevation of works on line A B C D E F of Fig. -42.] - -Figs. 43 to 46 show one of the furnaces. It will be seen at once that -its construction is similar to that of the standard American furnaces. -Pilz furnaces were tried in the first few years, but were finally -abandoned, as they could not be kept running for any satisfactory -length of time with slags rich in zinc. Diluting the slag, on the other -hand, would have led to an increased coke consumption, and would have -rendered the slag itself worthless. The furnace, however, differs in -several respects from its American prototype; the following are some of -the chief characteristics peculiar to it: - -[Illustration: Section E F. Section G H. -FIG. 44.—Shaft Furnace for Lead Smelting.] - -The chimney above the feed-floor covers one-third of the furnace -shaft, and is turned down in the form of a siphon, to connect with -the flue-dust chamber. The lateral faces, which are left open, serve -as charging apertures; the central one of these, provided with a -counterbalanced sheet-iron door, is used for charging from cars. The -square openings at the ends, which are covered with cast-iron plates, -are used for barring down the furnace shaft and may also be used for -charging. By this arrangement, together with the two hoppers placed -laterally on the chimney, it is possible to distribute the charge in -any desired manner over the whole cross-section of the furnace. This -arrangement greatly facilitates the removal of any accretions in the -furnace shaft, as the centrally placed chimney catches all the smoke, -while the charge-holes render the furnace accessible on all sides. -In case of large accretions being formed, the whole furnace can be -emptied, cleaned and restarted in 24 to 36 hours. - -The smelting cone is enclosed by cast-steel plates 50 cm. high, instead -of having a water-jacket. These are cooled as desired by turning a -jet of water on them. The plates are connected to the furnace shaft -by a bosh wall 25 cm. thick, which is surrounded with a boiler-plate -jacket. These jacket plates also are cooled from the outside by sprays -of water. With this arrangement the consumption of water is less than -with water-jackets, as a part of the water is vaporized, and the danger -of leakage of the jackets is avoided. The cast-steel plates are made -in two patterns; there are two similar side-plates, each with four -slits for the tuyeres, and two end-plates, provided with a circular -breast of 30 cm. aperture, for tapping the slag. The breast is cooled -by water flowing down, and is closed in front by a plate of sheet -iron, in which is the tap-hole for running off the slag. When cleaning -out, this sheet-iron plate is removed and the breast is opened, thus -providing easy access to the hearth. The four cast-steel plates are -anchored together with bolts at their outer ribs, and rest on two long, -gutter-shaped pieces of sheet iron, which carry off all the water which -flows down, and keep it away from the brickwork of the hearth. - -[Illustration: Section J L. Section C D. -FIG. 45.—Shaft Furnace.] - -The hearth, cased with boiler plate and rails, has at the side a -cast-iron pipe of 10 cm. diameter for drawing off the lead to the -outside kettle; this pipe has a slight downward inclination, to prevent -the slag flowing out; every 20 minutes lead is tapped, and the end of -the pipe is then plugged up with clay. - -The furnace shaft is supported upon a hollow mantel, which serves at -the same time as blast-pipe. The blast-pipe has eight lateral tees, -which are connected by canvas hose with the eight tuyeres. The mouth -of the tuyeres has the form of a horizontal slit, whereby the air is -distributed more evenly over the entire zone of fusion. - -[Illustration: FIG. 46.—Shaft Furnace for Lead Smelting. (Section A B.)] - -The precipitation of flue dust is effected in a brick condensing -chamber, placed near the beginning of the main flue. The main flue -terminates on the hill (see Fig. 43) in a chimney, the top of which -is 160 m. above the ground level of the works, affording excellent -draft. The condensing chamber (Figs. 49 to 51) consists of a vaulted -room, 3.40 m. wide and 6.60 m. long, which is divided into twelve -compartments by one longitudinal and five baffle walls. The gases -change direction seven times, and pass over the longitudinal wall -six times, being struck six times by fine sprays of water. The six -atomizers for this purpose consume 1.5 liter of water per minute, of -which four-fifths is vaporized, while one-fifth flows off to the lower -water basin. By this means 10 to 15 per cent. of the total flue dust -is precipitated in the condensing chamber itself, and is removed from -time to time as mud through the lower openings, which are water-sealed. -The remainder of the volatilized water precipitates the flue dust -almost completely on the way to the stack, so that only a short column -of steam is visible at the mouth of the stack. The flue to the stack -passes for the most part underground through abandoned adits and -galleries, thus providing a variety of changes in cross-section and -in direction, and assisting materially the action of the condensing -chamber. - -[Illustration: FIG. 47.—Section of Lead Refinery.] - -[Illustration: FIG. 48.—Softening Furnace.] - -As the charge of the shaft furnaces is poor in sulphur, no real matte -is produced, but only work lead and lead ashes (Bleischaum), which -contains 90 per cent. of lead, 1.6 per cent. sulphur, 0.4 per cent. -zinc, 0.85 per cent. Cu, 0.99 per cent. Fe, and 0.22 per cent. Sb. By -liquation and a reducing smelt in a reverberatory furnace, most of the -lead is obtained, along with a lead-copper matte, which is smelted for -copper matte and antimonial lead in the blast furnace. - -[Illustration: FIG. 49.—Fume Condenser. (Section A B.)] - -The copper matte, containing 18 per cent. Cu, 25 per cent. Fe, 30 per -cent. Pb and 18.4 per cent. S, is roasted dead in a reverberatory -furnace, is sintered, and melted to copper-bottoms in a small shaft -furnace. These copper-bottoms, which contain 60 per cent. copper and -25 per cent. lead, are subjected to liquation, and finally refined to -blister copper. - -The zinc-desilvering plant, Fig. 47, consists of a reverberatory -softening furnace, two desilvering kettles of 14 tons capacity, a pan -for liquating the zinc crust, and a small kettle for receiving the lead -from the liquation process. - -This pan has the advantage over the ordinary liquating kettle, that the -lead which drips off is immediately removed, before it can dissolve the -alloy; the silver content of the liquated lead is scarcely 0.05 per -cent., while the dry alloy contains 5 to 8 per cent. - -[Illustration: FIG. 50.—Fume Condenser. (Section E F G H.)] - -[Illustration: FIG. 51.—Fume Condenser. (Section C D.)] - -The removal of the zinc is effected in a second reverberatory furnace. -Formerly the steam-method was used, but the rapid wear of the kettles, -and the excessive formation of oxides called for a change in the -process. The zinc-silver alloy is distilled in a crucible of 200 kg. -capacity, and is cupeled in an English cupel furnace. The details of -the reverberatory furnace are shown in Fig. 48. - -The composition of the final products is shown by the following -analyses; Lead: Zn, 0.0021 per cent.; Fe, 0.0047 per cent.; Cu, 0.0005 -per cent.; Sb, 0.0030 per cent.; Bi, 0.0007 per cent.; Ag, 0.0010 per -cent.; Pb, 99.998 per cent.; Silver, Ag, 99.720 per cent.; Cu, 0.121 -per cent.; Fe, 0.005 per cent.; Pb, 0.018 per cent.; Au, 0.003 per -cent. - - - - -INDEX - - - Alloy, retorting the, in lead refining, 267 - - Alumina, experience with, 259 - - American Smelting and Refining Co., 4, 6, 26, 93, 113, 252, 295 - at Murray, Utah, 287 - - Atmosphere, effect of on concrete, 242 - - - Bag-house, cost of attending, 246 - standard, 246 - - Bag-houses for saving fume, 244 - - Bartlett, Eyre O., 244 - - Bayston, W. B., 199 - - Bennett, James C., 66 - - Betts, Anson G., 270, 274 - - Between products, working up of, 39 - - Biernbaum, A., 41, 148, 160 - - Blast furnace of circular form, 253 - Spanish lead, 307 - - Blast, volume and pressure of in lead smelting, 76 - - Blower, rotary, deficiency of, 251 - - Blowers for lead and copper smelting, 256 - now more powerful for lead smelting use, 252 - - Blowers, rotary, method of testing volumetric efficiency of, 254 - _vs._ blowing engines, 254 - _vs._ blowing engines for lead smelting, 251 - - Blowing engines, when to use, 259 - - Bonne Terre lead deposits, 18 - orebody, Missouri, 13, 14 - - Borchers, W., 114, 116, 127 - - Bormettes method, combination processes in, 222 - - Bradford, Mr., 55 - - Bretherton, S. E., 251, 258 - - Broken Hill Proprietary Block, 14, 59 - - Broken Hill practice, 51 - Proprietary Co., 52, 113, 124, 145, 175, 178, 206 - - Bricking plant for flue dust and fine ores, 66-70 - - Briquetting costs, 62 - methods of avoiding, 63, 64 - process, operations, in 59 - - Bullion, analyses of in lead refining, 281 - refined lead and slimes, analyses of, 282 - - - Canadian Smelting Works, 275 - - Carlton Iron Co., 63 - - Carmichael, A. D. 56, 199 - - Carmichael-Bradford process, 175-185 - brief estimate of, 209 - claims of in patent, 199 - recommendations of, 124 - process, points concerning, 131 - - Cement walls, how to build, 241 - - Channing, J. Parke, 254 - - Charge-car in smelting, true function of, 94 - feeding of in lead smelting, 77 - mechanical character of in lead smelting, 78 - - Charges, effect of large in lead smelting, 77 - - Cherokee Lanyon Smelter Co., 104 - - Chimney bases, 237 - - Chisholm, Boyd & White Co., 64 - - Clark, Donald, 114, 144, 175 - - Cœur d’Alene mines, 5, 6, 7 - - Concrete flues and stacks, advantages and disadvantages of, 242 - in metallurgical construction, 234 - - Connersville Blower Co., 252 - - Consolidated Kansas City Smelting and Refining Co., 285 - - Coke, percentage necessary to use in smelting, 259 - - Croll, H. V., 253 - - Cupellation in lead refining, 269 - - - De Lamar Copper Refining Co., 297 - - Desilverization in lead refining, 265 - - Desloge practice contrasted with others, 46 - - Doeltz, F. O., 139 - - Dross, analyses of in lead refining, 279 - - Dupuis & Sons, 63 - - Dust chamber, arched form, 231 - beehive form of, 232 - design, 229 - rectangular form, 230 - concrete, 235-237 - - Dwight, Arthur S., 73, 81 - spreader and curtain in furnaces, 91 - - - East Helena and Pueblo smelting systems compared, 93 - plant of the American Smelting and Refining Co., 302 - system of smelting, 88-94 - - Edwards, Henry W., 234, 240, 242 - - Einstein silver mine, 14 - - Engine, blowing, proper field of, 257 - blowing, and rotary blowers, 258 - - Eriksson, Hjalmar, 306 - - - Federal Lead Co., 38 - Mining and Smelting Co., 7 - - Feeders, cup and cone, for round furnaces, 81 - - Ferraris, Erminio, 311 - - Flat River mines, 18 - - Flue gases and moisture, effect of on concrete, 242 - - Flues, concrete, 234, 240, 242 - - Foundations for dynamos, 236 - - Fremantle Smelting Works, 145 - - Fume-smelting, cost of, 33 - in the hearth, 32 - - Furnace operations at Desloge, Mo., 45 - - Furnaces at Desloge, Mo., 43 - reverberatory, at Desloge, Mo., 42 - - - Galena, experiments in roasting, 129 - lime-roasting of, 14 - new methods of desulphurizing, 116 - roasting of by Savelsberg process, 122, 123 - - Gas, furnace, effect of on cement, 240 - - Gelatine, use of in electrolytic lead refining, 275 - - Germot, A., 224 - process, 224 - - Globe plant of the American Smelting and Refining Co., 304 - Smelting and Refining Co., 244 - - Greenway, T. J., 59 - - Guillemain, C., 133 - - - Harvard, Francis T., 242 - - Hearth, covered-in, 36 - - Heat, effect of on cement, 242 - - Heberlein, Ferdinand, 113, 167, 199 - - Hixon, Hiram W., 256, 258 - - Harwood, E. J., 51 - - Hourwich, Dr. Isaac A., 27 - - Huntington-Heberlein process, 113, 144-147 - consideration and estimate of, 203-209 - credit due to, 126 - process as distinguished from others, 118 - economic results of, 155-159 - - Huntington-Heberlein explained by the inventors, 167-173 - process at Friedrichshütte, 148 - process, from the hygienic standpoint, 160 - ideas of in patent specifications, 117 - process, introduction of at Tarnowitz, Prussia, 41 - and Savelsberg processes, essential difference between, 192 - process, some disadvantages of, 165, 166 - - Huppertz, L., 121 - - Hutchings, W. Maynard, 108, 126, 170 - - Huntington, Thomas, 113, 167, 199 - - - Iles, Malvern W., 96, 252 - - Ingalls, W. R., 3, 16, 27, 42, 177, 186, 193, 215, 224, 244, 287 - - Iron, behavior of in silver-lead smelting, 75 - - - Jackson Revel mine, 14 - - Johnson, E. M., 104 - R. D. O., 18 - - Jones, Richard, 244 - Samuel T., 244 - - - Laur, F., 224 - - Lead, analyses of refined, 281 - bullion, electrolytic refining of base, 270 - bullion, Parkes process of desilverizing and refining, 263 - bullion, softening of, 263 - concentrate Joplin district, valuation of, 25 - and copper smelting, the Bormettes method of, 215-223 - deposits, southeastern Missouri, 18 - Joplin district, 8 - marketing, 3 - -ore roasting, consideration of new processes, 135-138 - - Lead ore, average prices for, 27 - ore, cost of smelting, 32 - -ore roasting, theoretical aspects of, 133 - ores, Galena, Kan., 24 - ores, method of valuing, 26 - ores, southwestern Missouri, 24 - Park City, Utah, 8 - -poisoning in old and new processes, 162-165 - refining, electrolytic, 274 - soft, Missouri, 25 - smelting at Desloge, Mo., 42 - smelting at Monteponi, Sardinia, 311 - smelting and refining, cost of, 96 - smelting in the Scotch hearth, 31 - smelting in Spain, 306 - smelting at Tarnowitz, Prussia, 41 - source of in Missouri, 13 - in southeastern Missouri, 7, 10, 17 - sulphide and calcium sulphate, metallurgical behavior of, 139-143 - total production United States, 5 - yield from Scotch hearths, 39 - - Leadville, Colo., mines, 8 - - Lewis, G. T., 244 - - Lime-roasting of galena, 126 - - Lotti, Alfredo, 215 - - - Messiter, Edwin H., 229, 240 - - Middleton, K. W. M., 31 - - Mine La Motte, 14 - - Minerals, briquetting of, 63 - - Mining methods in Missouri, 19-23 - - Missouri Smelting Co., 197 - - Mould, H. S., Co., 64 - - Murray smelter, Utah, 291 - - - National plant of the American Smelting and Refining Co., 299 - - New Jersey Zinc Co., 246 - - Nutting, Mr., 256 - - - Ore and Fuel Co., 63 - different behavior of coarse and fine in lead smelting, 79 - treatment in detail by the Huntington-Heberlein process, 150-155 - - - Parkes process, cost of refining by, 99 - - Percy, Dr., 244 - - Perth Amboy plant of the American Smelting and Refining Co., 296 - - Petraeus, C. V., 24 - - Pfort curtain for furnaces, 82 - - Picher Lead Co., 197 - - Piddington, F. L., 263 - - Potter, Prof. W. B., 15 - - Pueblo lead smelter, 294 - - Smelting and Refining Co., 84 - - Pufahl, O., 38, 291, 294, 296, 299, 302, 304 - - Pyritic smelting without fuel practically impossible, 256 - - - Raht, August, 251, 254 - - Refining, monthly cost of per ton of bullion treated, 100 - - Roasters, hand, and mechanical furnaces, average monthly cost of, 98 - - Roberts-Austen, W. C., 139 - - - Salts, effect of crystallization of contained on concrete, 243 - - Santa Fe Gold and Copper Mining Co., 255 - - Savelsberg, Adolf, 122 - - Savelsberg process, 186-192 - process, claims of in patent, 201 - process contrasted with Huntington-Heberlein, 209 - process, difference between and Huntington-Heberlein, 197 - - Savelsberg process the simplest, 132 - - Scotch-hearth method, permanency of, 195 - - Scotch hearths, 34 - - Schneider, A. F., 81 - - Seattle Smelting and Refining Works, 273 - - Silver-lead blast furnaces, mechanical feeding of, 81 - blast furnace, proper conditions, 73 - smelting, details of practice, 73 - smelting, modern, 73 - - Slag-smelting costs, 34 - - Slime analysis at Broken Hill, 51 - - Slimes, analyses of in lead refining, 281 - desulphurization of by heap roasting, 51 - treatment of at Broken Hill, 53-55 - - Smelter, new, at El Paso, Texas, 285 - - Smelters’ pay, 32 - - Smelting, average cost of per ton, 98 - - Smelting Co. of Australia, 263 - costs, 48 - detailed costs of, 101, 102 - of galena ore, 38 - preparation of fine material for, 59 - - Solution, washing from slime, 277 - - Sticht, Mr., 256 - - St. Joseph Lead Co., 16 - - St. Louis Smelting and Refining Co., 81 - - Sulphide Corporation, 145 - - Sulphur dioxide, effect of on cement, 240 - - Sulphuric acid, making of at Broken Hill, 174 - - - Tasmanian Smelting Co., 145 - - Tennessee Copper Co., 255 - - Terhune, R. H., furnace gratings, 84 - - Thacher, Arthur, 14 - - - Ulke, Titus, 270 - - United Smelting and Refining Co., 88 - States Zinc Co., 295 - - - Vezin, H. A., 252 - - - Walls, retaining, 237 - - Walter, E. W., 260 - - Waring, W. Geo., 24 - - Welch, Max J., 229 - - Wetherill, Samuel, 244 - - Wheeler, H. A., 10 - - - Zinc, amount required in lead refining, 265, 266 - crusts, treatment of in lead refining, 267 - oxide in slags, 108 - retort residues, analysis of materials smelted and - bullion produced, 106 - retort residues, smelting, 104 - - -FOOTNOTES: - -[1] During 1905, antimonial lead commanded a premium of about 1c. per -lb. above desilverized, owing to the high price for antimony. - -[2] The figures for 1903 and 1904 have been added in the revision of -this article for this book. The production of lead in the United States -in 1903 was 276,694 tons; in 1904, it was 302,204 tons. - -[3] Ounces of silver to the ton of lead. - -[4] These figures are doubtful; they are probably too high. (See table -on p. 5). - -[5] The production of zinc ore in this district has now been commenced. - -[6] The manuscript of this article was dated Oct. 5, 1905. - -[7] Translated from _Zeit. f. Berg.-Hütten-und Salinenwesen_, LIII -(1905, p. 450). - -[8] This paper is published in pp. 148-166 of this book. - -[9] Abstract from _Transactions_ of the Australasian Institute of -Mining Engineers, Vol. IX, Part 1. - -[10] In the course of subsequent discussion Mr. Horwood stated that the -losses in roasting were 12½ per cent. in lead and probably about 5 per -cent. in silver. As compared to roasting in Ropp furnaces the loss in -lead was 5 to 6 per cent. greater, but the difference of loss in silver -was, he thought, not appreciable. Mr. Hibbard said that the Central -mine had obtained satisfactory results with masonry kilns.—EDITOR. - -[11] Abstract of portion of a paper presented at the Mexican meeting -of the American Institute of Mining Engineers, under the title “The -Mechanical Feeding of Silver-Lead Blast Furnaces.” _Transactions_, Vol. -XXXII, pp. 353-395. - -[12] Abstract of a paper (“The Mechanical Feeding of Silver-Lead Blast -Furnaces”) presented at the Mexican meeting of the American Institute -of Mining Engineers and published in the _Transactions_, Vol. XXXII. -For the first portion of this paper see the preceding article. - -[13] Abstract of a paper in _Western Chemist and Metallurgist_, I, VII, -Aug., 1905. - -[14] Much better work is being done at present, smelting the Western -zinc ores, and the residue contains about one-third of the above -figure, or 7.5 per cent. of zinc oxide. The high per cent. of ZnO left -in residue was mainly due to poor roasting. - -[15] There was also considerable coke used of an inferior grade, made -from Kansas coal. - -[16] Part of the ZnO in roasted matte came from being roasted in the -same furnace the zinc ore had been roasted in. - -[17] There was less residue on the charges during this month, which -accounts for the larger tonnage with a lower blast. - -[18] Translation of a paper read before the Naturwissenschaftlicher -Verein at Aachen, and published in _Metallurgie_, 1905, II, i, 1-6. - -[19] 35 to 40 cm. = 13.78 to 15.75 in. = 8 to 9.12 oz. per sq. in. - -[20] _Engineering and Mining Journal_, 1904, LXXVIII, p. 630; article -by Donald Clark; reprinted in this work, p. 144. - -[21] Owner of the patents.—EDITOR. - -[22] Abstract of a paper in _Metallurgie_, II, 18, Sept. 22, 1905, p. -433. - -[23] This method is described further on in this book. - -[24] Translated from _Metallurgie_, Vol. II, No. 19. - -[25] British patent, No. 17,580, Jan. 30, 1902, “Improved process for -desulphurizing sulphide ores.” - -[26] W. C. Roberts-Austen, “An Introduction to the Study of -Metallurgy,” London, 1902. - -[27] A. Lodin, _Comptes rendus_, 1895, CXX, 1164-1167; _Berg. u. -Hüttenm. Ztg._, 1903, p. 63. - -[28] _Comptes rendus_, loc. cit. - -[29] Translated from the _Zeitschrift für das Berg.-Hütten-und -Salinenwesen im. preuss. Staate_, 1905, LIII, ii, pp. 219-230. - -[30] Translated from the _Zeitschrift für das Berg.-Hütten-und -Salinenwesen im. preuss. Staate_, 1905, LIII, ii, pp. 219-230. - -[31] The manufacture of sulphuric acid from these gases has now been -undertaken in Silesia on a working scale.—EDITOR. - -[32] A paper presented before the American Institute of Mining -Engineers, July, 1906. - -[33] _Engineering and Mining Journal_, Sept. 2, 1905. - -[34] This term is inexact, because the hearths employed in the United -States are not strictly “Scotch hearths,” but they are commonly known -as such, wherefore my use of the term. - -[35] Percentages of lead in Missouri practice are based on the wet -assay; among the silver-lead smelters of the West the fire assay is -still generally employed. - -[36] This improvement did not originate at either Alton or -Collinsville. It had previously been in use at the works of the -Missouri Smelting Company at Cheltenham, St. Louis, but the idea -originated from the practice of the Picher Lead Company, of Joplin, Mo. - -[37] This refers especially to the Savelsberg process. - -[38] A. D. Carmichael, U. S. patent No. 705,904, July 29, 1902. - -[39] _Metallurgie_, 1905, II, i, 1-6; _Engineering and Mining Journal_, -Sept. 2, 1905. - -[40] _Metallurgie_, 1905, II, 19; _Engineering and Mining Journal_, -Jan. 27, 1906. - -[41] _Metallurgie_, 1905, Sept. 22, 1905; _Engineering and Mining -Journal_, March 10, 1906. - -[42] _Engineering and Mining Journal_, Oct. 21, 1905. - -[43] Translated by W. R. Ingalls. - -[44] As originally published the title of this article was -“Lead-Smelting without Fuel.” In this connection reference may well be -made to Hannay’s experiments and theories, _Transactions_ Institution -of Mining and Metallurgy, II, 188, and Huntington’s discussion, -_ibid._, p. 217. - -[45] Excerpt from a paper, “Concrete in Mining and Metallurgical -Engineering,” _Transactions_ American Institute of Mining Engineers, -XXXV (1905), p. 60. - -[46] A Discussion of the Paper by Henry W. Edwards, on “Concrete in -Mining and Metallurgical Engineering,” _Transactions_ of the American -Institute of Mining Engineers, XXXV. - -[47] _Engineering News_, Nov 30, 1899, and U. S. Patent No. 665,250, -Jan. 1 1901. - -[48] A discussion of the paper of Henry W. Edwards, on “Concrete in -Mining and Metallurgical Engineering,” _Transactions_ of the American -Institute of Mining Engineers, XXXV. - -[49] Abstract from the _Journal_ of the Chemical, Metallurgical and -Mining Society of South Africa, May, 1903. - -[50] Abstract of a paper in _Transactions_ American Institute of Mining -Engineers, XXXIV (1904), p. 175. - -[51] Silver not given. This was the case, also, with the gold in the -bullion. The slimes contained 0.131 per cent. of gold, or 39.1 oz. per -ton. - -[52] A constituent company of the American Smelting and Refining -Company. - -[53] Translated from _Zeit. f. Berg.-Hütten.-und Salinenwesen im -preuss. Staate_, 1905, LIII, p. 433. - -[54] Abstract from a paper in _Zeit. f. Berg.-Hütten-und Salinenwesen -im preuss. Staate_, 1905, LIII, p. 439. - -[55] Translated from _Zeit. f. Berg.-Hütten.-und Salinenwesen im -preuss. Staate_, 1905, LIII, 490. - -[56] Abstract from a paper in _Zeit. f. Berg.-Hütten-und Salinenwesen -im preuss. Staate_, 1905, p. 400. - -[57] Abstract from a paper in _Zeit. f. Berg.-Hütten.-und Salinenwesen -im preuss. Staate_, 1905, p. 400. - -[58] Abstract from an article in _Zeit. f. Berg.-Hütten.-und -Salinenwesen im preuss. Staate_, 1905, LIII, p. 444. - -[59] Translated from _Oest. Zeit. f. Berg.-und Hüttenwesen_, 1905, p. -455. - - - - - - -End of the Project Gutenberg EBook of Lead Smelting and Refining, by Various - -*** END OF THIS PROJECT GUTENBERG EBOOK LEAD SMELTING AND REFINING *** - -***** This file should be named 63784-0.txt or 63784-0.zip ***** -This and all associated files of various formats will be found in: - http://www.gutenberg.org/6/3/7/8/63784/ - -Produced by deaurider, Les Galloway and the Online -Distributed Proofreading Team at https://www.pgdp.net (This -file was produced from images generously made available -by The Internet Archive) - -Updated editions will replace the previous one--the old editions will -be renamed. - -Creating the works from print editions not protected by U.S. copyright -law means that no one owns a United States copyright in these works, -so the Foundation (and you!) can copy and distribute it in the United -States without permission and without paying copyright -royalties. 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