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 cover was prepared by the transcriber and is placed in the public domain.

LEAD SMELTING
AND
REFINING

WITH SOME NOTES ON LEAD MINING

EDITED BY
WALTER RENTON INGALLS

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

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]

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:

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:

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

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

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

FUME RECEIVING A PRELIMINARY ROAST

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:

Recent analyses of Weardale (Durham county) lead smelted in the Scotch hearth, and slag-lead smelted in the blast furnace, are given herewith:

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:

With three shifts on a furnace, the schedule was as follows:

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.

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:

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, 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.