CONTENTS
| PART I | |
| Notes on Lead Mining | |
| PAGE | |
| Sources of Lead Production in the United States (WalterRenton 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 andW. 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 BrokenHill (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 andFerdinand 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 (AlfredoLotti) | [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 atMurray, 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 RefiningCompany (O. Pufahl) | [296] |
| The National Plant of the American Smelting and RefiningCompany (O. Pufahl) | [299] |
| The East Helena Plant of the American Smelting and RefiningCompany (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 | i>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 |
| Smelters’ 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 |
| Roasters’ coal (425 lb.) | 0.65 |
| Smelters’ wages | 1.08 |
| Smelters’ 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 |
| Smelters’ coal (1400 lb.) | 2.13 |
| Total | $5.01 |
FUME RECEIVING A PRELIMINARY ROAST
| Roasters’ wages | $2.08 |
| Roasters’ coal (1450 lb.) | 2.18 |
| Smelters’ wages | 2.04 |
| Smelters’ 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 Hearth | Silver-Lead from Hearth | Slag-Lead from Blast Furnace | |
|---|---|---|---|
| Lead | 99.957 | 99.957 | 99.013 |
| Silver | 0.0035 | 0.0200 | 0.0142 |
| (1 oz. 2 dwt. 21 gr. per Long Ton) | (6 oz. 10 dwt. 16 gr. per Long Ton) | (4 oz. 12 dwt. 18 gr. 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.