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Title: Lead Smelting and Refining - With notes on lead mining
Author: Various
Language: English
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Transcriber’s Notes
Obvious typographical errors have been silently corrected. Variations
in hyphenation other spelling and punctuation remains unchanged. In
particular the words height and hight are used about equally. As hight
is a legitimate spelling, it has not been changed.
Some of the larger tables have been re-organised to improve clarity and
avoid excessive width.
The footnotes are located at the end of the book.
Italics are represented thus _italic_.
LEAD SMELTING
AND
REFINING
WITH SOME NOTES ON LEAD MINING
EDITED BY
WALTER RENTON INGALLS
[Illustration: Publisher’s Device]
NEW YORK AND LONDON
THE ENGINEERING AND MINING JOURNAL
1906
COPYRIGHT, 1906,
BY THE ENGINEERING AND MINING JOURNAL.
ALSO ENTERED AT
STATIONERS’ HALL, LONDON, ENGLAND.
ALL RIGHTS RESERVED.
PREFACE
This book is a reprint of various articles pertaining especially to the
smelting and refining of lead, together with a few articles relating
to the mining of lead ore, which have appeared in the _Engineering and
Mining Journal_, chiefly during the last three years; in a few cases
articles from earlier issues have been inserted, in view of their
special importance in rounding out certain of the subjects treated.
For the same reason, several articles from the _Transactions_ of
the American Institute of Mining Engineers have been incorporated,
permission to republish them in this way having been courteously
granted by the Secretary of the Institute. Certain of the other
articles comprised in this book are abstracts of papers originally
presented before engineering societies, or published in other technical
periodicals, subsequently republished in the _Engineering and Mining
Journal_, as to which proper acknowledgment has been made in all cases.
The articles comprised in this book relate to a variety of subjects,
which are of importance in the practical metallurgy of lead, and
especially in connection with the desulphurization of galena, which is
now accomplished by a new class of processes known as “Lime Roasting”
processes. The successful introduction of these processes into the
metallurgy of lead has been one of the most important features in
the history of the latter during the last twenty-five years. Their
development is so recent that they are not elsewhere treated in
technical literature, outside of the pages of the periodicals and the
transactions of engineering societies. The theory and practice of these
processes are not yet by any means well understood, and a year or two
hence we shall doubtless possess much more knowledge concerning them
than we have now. Prompt information respecting such new developments
is, however, more desirable than delay with a view to saying the
last word on the subject, which never can be said by any of us, even
if we should wait to the end of the lifetime. For this reason it
has appeared useful to collect and republish in convenient form the
articles of this character which have appeared during the last few
years.
W. R. INGALLS.
AUGUST 1, 1906.
CONTENTS
PART I
NOTES ON LEAD MINING
PAGE
SOURCES OF LEAD PRODUCTION IN THE UNITED STATES (WALTER
RENTON INGALLS) 3
NOTES ON THE SOURCE OF THE SOUTHEAST MISSOURI LEAD (H. A.
WHEELER) 10
MINING IN SOUTHEASTERN MISSOURI (WALTER RENTON INGALLS) 16
LEAD MINING IN SOUTHEASTERN MISSOURI (R. D. O. JOHNSON) 18
THE LEAD ORES OF SOUTHWESTERN MISSOURI (C. V. PETRAEUS AND
W. GEO. WARING) 24
PART II
ROAST-REACTION SMELTING
SCOTCH HEARTHS AND REVERBERATORY FURNACES
LEAD SMELTING IN THE SCOTCH HEARTH (KENNETH W. M. MIDDLETON) 31
THE FEDERAL SMELTING WORKS, NEAR ALTON, ILL. (O. PUFAHL) 38
LEAD SMELTING AT TARNOWITZ (EDITORIAL) 41
LEAD SMELTING IN REVERBERATORY FURNACES AT DESLOGE, MO.
(WALTER RENTON INGALLS) 42
PART III
SINTERING AND BRIQUETTING
THE DESULPHURIZATION OF SLIMES BY HEAP ROASTING AT BROKEN
HILL (E. J. HORWOOD) 51
THE PREPARATION OF FINE MATERIAL FOR SMELTING (T. J. GREENWAY) 59
THE BRIQUETTING OF MINERALS (ROBERT SCHORR) 63
A BRICKING PLANT FOR FLUE DUST AND FINE ORES (JAS. C. BENNETT) 66
PART IV
SMELTING IN THE BLAST FURNACE
MODERN SILVER-LEAD SMELTING (ARTHUR S. DWIGHT) 73
MECHANICAL FEEDING OF SILVER-LEAD BLAST FURNACES (ARTHUR S.
DWIGHT) 81
COST OF SMELTING AND REFINING (MALVERN W. ILES) 96
SMELTING ZINC RETORT RESIDUES (E. M. JOHNSON) 104
ZINC OXIDE IN SLAGS (W. MAYNARD HUTCHINGS) 108
PART V
LIME-ROASTING OF GALENA
THE HUNTINGTON-HEBERLEIN PROCESS 113
LIME-ROASTING OF GALENA (EDITORIAL) 114
THE NEW METHODS OF DESULPHURIZING GALENA (W. BORCHERS) 116
LIME-ROASTING OF GALENA (W. MAYNARD HUTCHINGS) 126
THEORETICAL ASPECTS OF LEAD-ORE ROASTING (C. GUILLEMAIN) 133
METALLURGICAL BEHAVIOR OF LEAD SULPHIDE AND CALCIUM SULPHATE
(F. O. DOELTZ) 139
THE HUNTINGTON-HEBERLEIN PROCESS (DONALD CLARK) 144
THE HUNTINGTON-HEBERLEIN PROCESS AT FRIEDRICHSHÜTTE (A.
BIERNBAUM) 148
THE HUNTINGTON-HEBERLEIN PROCESS FROM THE HYGIENIC STANDPOINT
(A. BIERNBAUM) 160
THE HUNTINGTON-HEBERLEIN PROCESS (THOMAS HUNTINGTON AND
FERDINAND HEBERLEIN) 167
MAKING SULPHURIC ACID AT BROKEN HILL (EDITORIAL) 174
THE CARMICHAEL-BRADFORD PROCESS (DONALD CLARK) 175
THE CARMICHAEL-BRADFORD PROCESS (WALTER RENTON INGALLS) 177
THE SAVELSBERG PROCESS (WALTER RENTON INGALLS) 186
LIME-ROASTING OF GALENA (WALTER RENTON INGALLS) 193
PART VI
OTHER METHODS OF SMELTING
THE BORMETTES METHOD OF LEAD AND COPPER SMELTING (ALFREDO
LOTTI) 215
THE GERMOT PROCESS (WALTER RENTON INGALLS) 224
PART VII
DUST AND FUME RECOVERY
FLUES, CHAMBERS AND BAG-HOUSES
DUST CHAMBER DESIGN (MAX J. WELCH) 229
CONCRETE IN METALLURGICAL CONSTRUCTION (HENRY W. EDWARDS) 234
CONCRETE FLUES (EDWIN H. MESSITER) 240
CONCRETE FLUES (FRANCIS T. HAVARD) 242
BAG-HOUSES FOR SAVING FUME (WALTER RENTON INGALLS) 244
PART VIII
BLOWERS AND BLOWING ENGINES
ROTARY BLOWERS VS. BLOWING ENGINES FOR LEAD SMELTING (EDITORIAL) 251
ROTARY BLOWERS VS. BLOWING ENGINES (J. PARKE CHANNING) 254
BLOWERS AND BLOWING ENGINES FOR LEAD AND COPPER SMELTING
(HIRAM W. HIXON) 256
BLOWING ENGINES AND ROTARY BLOWERS (S. E. BRETHERTON) 258
PART IX
LEAD REFINING
THE REFINING OF LEAD BULLION (F. L. PIDDINGTON) 263
THE ELECTROLYTIC REFINING OF BASE LEAD BULLION (TITUS ULKE) 270
ELECTROLYTIC LEAD REFINING (ANSON G. BETTS) 274
PART X
SMELTING WORKS AND REFINERIES
THE NEW SMELTER AT EL PASO, TEXAS (EDITORIAL) 285
NEW PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY AT
MURRAY, UTAH (WALTER RENTON INGALLS) 287
THE MURRAY SMELTER, UTAH (O. PUFAHL) 291
THE PUEBLO LEAD SMELTERS (O. PUFAHL) 294
THE PERTH AMBOY PLANT OF THE AMERICAN SMELTING AND REFINING
COMPANY (O. PUFAHL) 296
THE NATIONAL PLANT OF THE AMERICAN SMELTING AND REFINING
COMPANY (O. PUFAHL) 299
THE EAST HELENA PLANT OF THE AMERICAN SMELTING AND REFINING
COMPANY (O. PUFAHL) 302
THE GLOBE PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY
(O. PUFAHL) 304
LEAD SMELTING IN SPAIN (HJALMAR ERIKSSON) 306
LEAD SMELTING AT MONTEPONI, SARDINIA (ERMINIO FERRARIS) 311
PART I
NOTES ON LEAD MINING
SOURCES OF LEAD PRODUCTION IN THE UNITED STATES
BY WALTER RENTON INGALLS
(November 28, 1903)
Statistics of lead production are of value in two directions: (1) in
showing the relative proportion of the kinds of lead produced; and (2)
in showing the sources from which produced. Lead is marketed in three
principal forms: (_a_) desilverized; (_b_) soft; (_c_) antimonial, or
hard. The terms to distinguish between classes “a” and “b” are inexact,
because, of course, desilverized lead is soft lead. Desilverized lead
itself is classified as “corroding,” which is the highest grade, and
ordinary “desilverized.” Soft lead, referring to the Missouri product,
may be either “ordinary” or “chemical hard.” The latter is such lead
as contains a small percentage of copper and antimony as impurities,
which, without making it really hard, increase its resistance against
the action of acids, and therefore render it especially suitable
for the production of sheet to be used in sulphuric-acid chamber
construction and like purposes. The production of chemical hard lead
is a fortuitous matter, depending on the presence of the valuable
impurities in the virgin ores. If present, these impurities go into
the lead, and cannot be completely removed by the simple process of
refining which is practised. Nobody knows just what proportions of
copper and antimony are required to impart the desired property, and
consequently no specifications are made. Some chemical engineers call
for a particular brand, but this is really only a whim, since the same
brand will not be uniformly the same; practically one brand is as good
as another. Corroding lead is the very pure metal, which is suitable
for white lead manufacture. It may be made either from desilverized or
from the ordinary Missouri product; or the latter, if especially pure,
may be classed as corroding without further refining. Antimonial lead
is really an alloy of lead with about 15 to 30 per cent. antimony,
which is produced as a by-product by the desilverizers of base
bullion. The antimony content is variable, it being possible for the
smelter to run the percentage up to 60. Formerly it was the general
custom to make antimonial lead with a content of 10 to 12 per cent. Sb;
later, with 18 to 20 per cent.; while now 25 to 30 per cent. Sb is best
suited to the market.
The relative values of the various grades of lead fluctuate
considerably, according to the market place, and the demand and supply.
The schedules of the American Smelting and Refining Company make a
regular differential of 10c. per 100 lb. between corroding lead and
desilverized lead in all markets. In the St. Louis market, desilverized
lead used to command a premium of 5c. to 10c. per 100 lb. over ordinary
Missouri; but now they sell on approximately equal terms. Chemical hard
lead sells sometimes at a higher price, sometimes at a lower price,
than ordinary Missouri lead, according to the demand and supply. There
is no regular differential. This is also the case with antimonial
lead.[1]
The total production of lead from ores mined in the United States in
1901 was 279,922 short tons, of which 211,368 tons were desilverized,
57,898 soft (meaning lead from Missouri and adjacent States) and
10,656 antimonial. These are the statistics of “The Mineral Industry.”
The United States Geological Survey reported substantially the same
quantities. In 1902 the production was 199,615 tons of desilverized,
70,424 tons of soft, and 10,485 tons of antimonial, a total of 280,524
tons. There is an annual production of 4000 to 5000 tons of white
lead direct from ore at Joplin, Mo., which increases the total lead
production of the United States by, say, 3500 tons per annum. The
production of lead reported as “soft” does not represent the full
output of Missouri and adjacent States, because a good deal of their
ore, itself non-argentiferous, except to the extent of about 1 oz. per
ton in certain districts, is smelted with silver-bearing ores, going
thus into an argentiferous lead; while in one case, at least, the
almost non-argentiferous lead, obtained by smelting the ore unmixed, is
desilverized for the sake of the extra refining.
Lead-bearing ores are of widespread occurrence in the United States.
Throughout the Rocky Mountains there are numerous districts in which
the ore carries more or less lead in connection with gold and silver.
For this reason, the lead mining industry is not commonly thought of as
having such a concentrated character as copper mining and zinc mining.
It is the fact, however, that upward of 70 per cent. of the lead
produced in the United States is derived from five districts, and in
the three leading districts from a comparatively small number of mines.
The statistics of these for 1901 to 1904 are as follows:[2]
┌──────────┬───────────────────────────────┬───────────────────────┬────
│ │ PRODUCTION, TONS │ PER CENT. │
│DISTRICT │ 1901 │ 1902 │ 1903 │ 1904 │ 1901│ 1902│ 1903│ 1904│REF.
├──────────┼───────┼───────┼───────┼───────┼─────┼─────┼─────┼─────┼────
│Cœur │ │ │ │ │ │ │ │ │
│d’Alene │ 68,953│ 74,739│ 89,880│ 98,240│ 24.3│ 26.3│ 32.5│ 32.5│_a_
│Southeast │ │ │ │ │ │ │ │ │
│Mo. │ 46,657│ 56,550│ 59,660│ 59,104│ 16.4│ 19.9│ 21.2│ 19.6│_b_
│Leadville,│ │ │ │ │ │ │ │ │
│Colo. │ 28,180│ 19,725│ 18,177│ 23,590│ 10.0│ 6.9│ 6.6│ 7.8│_c_
│Park City,│ │ │ │ │ │ │ │ │
│Utah │ 28,310│ 36,300│ 36,534│ 30,192│ 10.0│ 12.8│ 13.2│ 10.0│_d_
│Joplin, │ │ │ │ │ │ │ │ │
│Mo.-Kan. │ 24,500│ 22,130│ 20,000│ 23,600│ 8.6│ 7.8│ 7.2│ 7.8│_e_
│ ├───────┼───────┼───────┼───────┼─────┼─────┼─────┼─────┤
│ Total │196,600│209,444│224,251│234,726│ 69.3│ 73.7│ 81.0│ 77.7│
_a._ The production in 1901 and 1902 is computed from direct returns
from the mines, with an allowance of 6 per cent. for loss of lead in
smelting. The production in 1903 and 1904 is estimated at 95 per cent.
of the total lead product of Idaho.
_b._ This figure includes only the output of the mines at Bonne Terre,
Flat River, Doe Run, Mine la Motte and Fredericktown. It is computed
from the report of the State Lead and Zinc Mine Inspector as to ore
produced, the ore (concentrates) of the mines at Bonne Terre, Flat
River and Doe Run being reckoned as yielding 60 per cent. lead.
_c._ Report of State Commissioner of Mines.
_d._ Report of Director of the Mint on “Production of Gold and Silver
in the United States,” with allowance of 6 per cent. for loss of lead
in smelting.
_e._ From statistics reported by “The Mineral Industry,” reckoning the
ore (concentrates) as yielding 70 per cent. lead.
Outside of these five districts, the most of the lead produced in the
United States is derived from other camps in Idaho, Colorado, Missouri
and Utah, the combined output of all other States being insignificant.
It is interesting to examine the conditions under which lead is
produced in the five principal districts.
_Leadville, Colo._—The mines of Leadville, which once were the largest
lead producers of the United States, became comparatively unimportant
after the exhaustion of the deposits of carbonate ore, but have
attained a new importance since the successful introduction of means
for separating the mixed sulphide ore, which occurs there in very large
bodies. The lead production of Leadville in 1897 was 11,850 tons;
17,973 tons in 1898; 24,299 tons in 1899; 31,300 tons in 1900; 28,180
tons in 1901, and 19,725 tons in 1902. The Leadville mixed sulphide ore
assays about 8 per cent. Pb, 25 per cent. Zn and 10 oz. silver per ton.
It is separated into a zinc product assaying about 38 per cent. Zn and
6 per cent. Pb, and a galena product assaying about 45 per cent. Pb, 10
or 12 per cent. Zn, and 10 oz. silver per ton.
_Cœur d’Alene._—The mines of this district are opened on fissure veins
of great extent. The ore is of low grade and requires concentration. As
mined, it contains about 10 per cent. lead and a variable proportion
of silver. It is marketed as mineral, averaging about 50 per cent. Pb
and 30 oz. silver per ton. The production of lead ore in this district
is carried on under the disadvantages of remoteness from the principal
markets for pig lead, high-priced labor, and comparatively expensive
supplies. It enjoys the advantages of large orebodies of comparatively
high grade in lead, and an important silver content, and in many cases
cheap water power, and the ability to work the mines through adit
levels. The cost of mining and milling a ton of crude ore is $2.50 to
$3.50. The mills are situated, generally, at some distance from the
mines, the ore being transported by railway at a cost of 8 to 20c.
per ton. The dressing is done in large mills at a cost of 40 to 50c.
per ton. About 75 per cent. of the lead of the ore is recovered. The
concentrates are sold at 90 per cent. of their lead contents and 95 per
cent. of their silver contents, less a smelting charge of $8 per ton,
and a freight rate of $8 per ton on ore of less than $50 value per ton,
$10 on ore worth $50 to $65, and $12 on ore worth more than $65; the
ore values being computed f. o. b. mines. The settling price of lead is
the arbitrary one made by the American Smelting and Refining Company.
With lead (in ore) at 3.5c. and silver at 50c., the value, f. o. b.
mines, of a ton of ore containing 50 per cent. Pb and 30 oz. silver is
approximately as follows:
1000 × 0.90 = 900 lb. lead, at 3.5c. $31.50
30 × 0.95 = 28.5 oz. silver, at 50c. 14.25
——————
Gross value, f. o. b. mines $45.75
Less freight, $10, and smelting charge, $8 18.00
——————
Net value, f. o. b. mines $27.75
Assuming an average of 6 tons of crude ore to 1 ton of concentrate, the
value per ton of crude ore would be about $4.62½, and the net profit
per ton about $1.62½, which figures are increased 23.75c. for each 5c.
rise in the value of silver above 50c. per ounce.
The production of the Cœur d’Alene since 1895, as reported by the
mines, has been as follows:
─—─—─—-———─—─┬—─—-——————─—─┬—─—-—————————┬——————─—─
YEAR │ LEAD, TONS │ SILVER, OZ. │ RATIO[3]
─—─—─—-———─—─┼—─—-——————─—─┼—─—-—————————┼——————─—─
1896 │ 37,250 │ 2,500,000 │ 67.1
1897 │ 57,777 │ 3,579,424 │ 61.9
1898 │ 56,339 │ 3,399,524 │ 60.3
1899 │ 50,006 │ 2,736,872 │ 54.7
1900 │ 81,535 │ 4,755,877 │ 58.3
1901 │ 68,953 │ 3,349,533 │ 48.5
1902 │ 74,739 │ 4,489,549 │ 60.0
1903 │ [4]100,355 │ 5,751,613 │ 57.3
1904 │ [4]108,954 │ 6,247,795 │ 57.4
─—─—─—-———─—─┴—─—-——————─—─┴—─—-—————————┴——————─—─
The number of producers in the Cœur d’Alene district is comparatively
small, and many of them have recently consolidated, under the name of
the Federal Mining and Smelting Company. Outside of that concern are
the Bunker Hill & Sullivan, the Morning and the Hercules mines, control
of which has lately been secured by the American Smelting and Refining
Company.
_Southeastern Missouri._—The most of the lead produced in this region
comes from what is called the disseminated district, comprising
the mines of Bonne Terre, Flat River, Doe Run, Mine la Motte and
Fredericktown, of which those of Bonne Terre and Flat River are the
most important. The ore of this region is a magnesian limestone
impregnated with galena. The deposits lie nearly flat and are very
large. They yield about 5 per cent. of mineral, which assays about 65
per cent. lead. The low grade of the ore is the only disadvantage which
this district has, but this is so much more than offset by the numerous
advantages, that mining is conducted very profitably, and it is an open
question whether lead can be produced more cheaply here or in the Cœur
d’Alene. The mines of southeastern Missouri are only 60 to 100 miles
distant from St. Louis, and are in close proximity to the coalfields
of southern Illinois, which afford cheap fuel. The ore lies at depths
of only 100 to 500 ft. below the surface. The ground stands admirably,
without any timbering. Labor and supplies are comparatively cheap.
Mining and milling can be done for $1.05 to $1.25 per ton of crude ore,
when conducted on the large scale that is common in this district.
Most of the mining companies are equipped to smelt their own ore, the
smelters being either at the mines or near St. Louis. The freight rate
on concentrates to St. Louis is $1.40 per ton; on pig lead it is $2.10
per ton. The total cost of producing pig lead, delivered at St. Louis,
is about 2.25c. per pound, not allowing for interest on the investment,
amortization, etc.
The production of the mines in the disseminated district in 1901 was
equivalent to 46,657 tons of pig lead; in 1902 it was 56,550 tons. The
milling capacity of the district is about 6000 tons per day, which
corresponds to a capacity for the production of about 57,000 tons of
pig lead per annum. The St. Joseph Lead Company is building a new 1000
ton mill, and the St. Louis Smelting and Refining Company (National
Lead Company) is further increasing its output, which improvements will
increase the daily milling capacity by about 1400 tons, and will enable
the district to put out upward of 66,000 tons of pig lead. In this
district, as in the Cœur d’Alene, the industry is closely concentrated,
there being only nine producers, all told.
_Park City, Utah._—Nearly all the lead produced by this camp comes
from the Silver King, Daly West, Ontario, Quincy, Anchor and Daly
mines, which have large veins of low-grade ore carrying argentiferous
galena and blende, a galena product being obtained by dressing, and
zinkiferous tailings, which are accumulated for further treatment as
zinc ore, when market conditions justify.[5]
_Joplin District._—The lead production of southwestern Missouri and
southeastern Kansas, in what is known as the Joplin district, is
derived entirely as a by-product in dressing the zinc ore of that
district. It is obtained as a product assaying about 77 per cent. Pb,
and is the highest grade of lead ore produced, in large quantity,
anywhere in the United States. It is smelted partly for the production
of pig lead, and partly for the direct manufacture of white lead. The
lead ore production of the district was 31,294 tons in 1895, 26,927
tons in 1896, 29,578 tons in 1897, 26,457 tons in 1898, 24,100 tons
in 1899, 28,500 tons in 1900, 35,000 tons in 1901, and 31,615 tons in
1902. The production of lead ore in this district varies more or less,
according to the production of zinc ore, and is not likely to increase
materially over the figure attained in 1901.
NOTES ON THE SOURCE OF THE SOUTHEAST MISSOURI LEAD
BY H. A. WHEELER
(March 31, 1904)
The source of the lead that is being mined in large quantities in
southeastern Missouri has been a mooted question. Nor is the origin
of the lead a purely theoretical question, as it has an important
bearing on the possible extension of the orebodies into the underlying
sandstone.
The disseminated lead ores of Missouri occur in a shaly, magnesian
limestone of Cambrian age in St. François, Madison and Washington
counties, from 60 to 130 miles south of St. Louis. The limestone
is known as the Bonne Terre, or lower half of “the third magnesian
limestone” of the Missouri Geological Survey, and rests on a sandstone,
known as “the third sandstone,” that is the base of the sedimentary
formations in the area. Under this sandstone occur the crystalline
porphyries and granites of Algonkian and Archean age, which outcrop as
knobs and islands of limited extent amid the unaltered Cambrian and
Lower Silurian sediments.
The lead occurs as irregular granules of galena scattered through the
limestone in essentially horizontal bodies that vary from 5 to 100
ft. in thickness, from 25 to 500 ft. in width, and have exceeded 9000
ft. in length. There is no vein structure, no crushing or brecciation
of the inclosing rock, yet these orebodies have well defined axes or
courses, and remarkable reliability and persistency. It is true that
the limestone is usually darker, more porous, and more apt to have thin
seams of very dark (organic) shales where it is ore-bearing than in the
surrounding barren ground. The orebodies, however, fade out gradually,
with no sharp line between the pay-rock and the non-paying, and the
lead is rarely, if ever, entirely absent in any extent of the limestone
of the region. While the main course of the orebodies seems to be
intimately connected with the axes of the gentle anticlinal folds,
numerous cross-runs of ore that are associated with slight faults are
almost as important as the main shoots, and have been followed for
5000 ft. in length. These cross-runs are sometimes richer than the
main runs, at least near the intersections, but they are narrower, and
partake more of the type of vertical shoots, as distinguished from the
horizontal sheet-form.
Most of the orebodies occur at, or close to, the base of the limestone,
and frequently in the transition rock between the underlying sandstone
and the limestone, though some notable and important bodies have
been found from 100 to 200 ft. above the sandstone. This makes the
working depth from the surface vary from 150 to 250 ft., for the upper
orebodies, to 300 to 500 ft. deep to the main or basal orebodies,
according as erosion has removed the ore-bearing limestone. The
thickness of the latter ranges from 400 to 500 ft.
Associated with the galena are less amounts of pyrite, which especially
fringes the orebodies, and very small quantities of chalcopyrite, zinc
blende, and siegenite (the double sulphide of nickel and cobalt).
Calcite also occurs, especially where recent leaching has opened
vugs, caves, or channels in the limestone, when secondary enrichment
frequently incrusts these openings with crystals of calcite and galena.
No barite ever occurs with the disseminated ore, though it is the
principal gangue mineral in the upper or Potosi member of the third
magnesian limestone, and is never absent in the small ore occurrences
in the still higher second magnesian limestone.
While the average tenor of the ore is low, the yield being from 3 to
4 per cent. in pig lead, they are so persistent and easy to mine that
the district today is producing about 70,000 tons of pig lead annually,
and at a very satisfactory profit. As the output was about 2500 tons
lead in 1873, approximately 8500 tons in 1883, and about 20,000 tons in
1893, it shows that this district is young, for the principal growth
has been within the last five years.
Of the numerous but much smaller occurrences of lead elsewhere in
Missouri and the Mississippi valley, none resembles this district
in character, a fact which is unique. For while the Mechernich lead
deposits, in Germany, are disseminated, and of even lower grade than in
Missouri, they occur in a sandstone, and (like all the lead deposits
outside of the Mississippi valley) they are argentiferous, at least to
an extent sufficient to make the extraction of the silver profitable;
and on the non-argentiferous character of the disseminated deposits
hangs my story.
Of the numerous hypotheses advanced to account for the origin of these
deposits, there are only two that seem worthy of consideration: (a)
the _lateral secretion theory_, and (b) _deposition from solutions of
deep-seated origin_. Other theories evolved in the pioneer period of
economic geology are interesting, chiefly by reason of the difficulties
under which the early strugglers after geological knowledge blazed the
pathway for modern research and observation.
The lateral secretion theory, as now modernized into the secondary
enrichment hypothesis, has much merit when applied to the southeastern
and central Missouri lead deposits. For the limestones throughout
Missouri—and they are the outcropping formation over more than half of
the State—are rarely, if ever, devoid of at least slight amounts of
lead and zinc, although they range in age from the Carboniferous down
to the Cambrian.
The sub-Carboniferous formation is almost entirely made up of
limestones, which aggregate 1200 to 1500 ft. in thickness. They
frequently contain enough lead (and less often zinc) to arouse the
hopes of the farmer, and more or less prospecting has been carried on
from Hannibal to St. Louis, or 125 miles along the Mississippi front,
and west to the central part of the State, but with most discouraging
results.
In the rock quarries of St. Louis, immediately under the lower coal
measures, fine specimens of millerite of world-wide reputation occur
as filiform linings of vugs in this sub-Carboniferous limestone. These
vugs occur in a solid, unaltered rock which gives no clue to the
existence of the vug or cavity until it is accidentally broken. The
vugs are lined with crystals of pink dolomite, calcite and millerite,
with occasionally barite, selenite, galena and blende. They occur
in a well-defined horizon about 5 ft. thick, and the vugs in the
limestone above and below this millerite bed contain only calcite,
or less frequently dolomite. Yet this sub-Carboniferous formation in
southwestern Missouri, about Joplin, carries the innumerable pockets
and sheets of lead and zinc that have made that district the most
important zinc producer in the world. While faulting and limited
folding occur in eastern and central Missouri to fully as great an
extent as in St. François county or the Joplin district, thus far no
mineral concentration into workable orebodies has been found in this
formation, except in the Joplin area.
The next important series of limestones that make up most of the
central portion of Missouri are of Silurian age, and in them lead and
zinc are liberally scattered over large areas. In the residual surface
clays left by dissolution of the limestone, the farmers frequently make
low wages by gophering after the liberated lead, and the aggregate
of these numerous though insignificant gopher-holes makes quite a
respectable total. But they are only worked when there is nothing else
to do on the farm, as with rare exceptions they do not yield living
wages, and the financial results of mining the rock are even less
satisfactory. Yet a few small orebodies have been found that were
undoubtedly formed by local leaching and re-precipitation of this
diffused lead and zinc. Such orebodies occur in openings or caves,
with well crystallized forms of galena and blende, and invariably
associated with crystallized “tiff” or barite. I am not aware of any
of these pockets or secondary enrichments having produced as much as
2000 tons of lead or zinc, and very few have produced as much as 500
tons, although one of these pockets was recently exploited with heroic
quantities of printer’s ink as the largest lead mine in the world. Yet
there are large areas in which it is almost impossible to put down a
drill-hole without finding “shines” or trifling amounts of lead or
zinc. That these central Missouri lead deposits are due to lateral
secretion there seems little doubt, and it is possible that larger
pockets may yet be found where more favorable conditions occur.
When the lateral secretion theory is applied to the disseminated
deposits of southeastern Missouri, we are confronted by enormous bodies
of ore, absence of barite, non-crystallized condition of the galena
except in local, small, evidently secondary deposits, and well-defined
courses for the main and cross-runs of ore. The Bonne Terre orebody,
which has been worked longest and most energetically, has attained a
length of nearly 9000 ft., with a production of about 350,000 tons
or $30,000,000 of lead, and is far from being exhausted. Orebodies
recently opened are quite as promising. The country rock is not as
broken nor as open as in central Missouri, and is therefore much less
favorable for the lateral circulation of mineral waters, yet the
orebodies vastly exceed those of the central region.
Further, the Bonne Terre formation is heavily intercalated with thick
sheets of shale that would hinder overlying waters from reaching the
base of the ore-horizon, where most of the ore occurs, so that the
leachable area would be confined to a very limited vertical range,
or to but little greater thickness than the 100 ft. or so in which
most of the orebodies occur. While I have always felt that such large
bodies, showing relatively rapid precipitation of the lead, could not
be satisfactorily explained except as having a deep-seated origin,
the fact that the disseminated ore is practically non-argentiferous,
or at least carries only one to three ounces per ton, has been a
formidable obstacle. For the lead in the small fissure-veins that
occasionally occur in the adjacent granite has always been reported
as argentiferous. Thus the Einstein silver mine, near Fredericktown,
worked a fissure-vein from 1 to 6 ft. wide in the granite. It had a
typical complex vein-filling and structure, and carried galena that
assayed from 40 to 200 oz. per ton. While the quantity of ore obtained
did not justify the expensive plant erected to operate it, the galena
was rich in silver, whereas in the disseminated ores at the Mine la
Motte mine, ten miles distant, only the customary 1.5 oz. per ton
occurs. Occasionally fine-grained specimens of galena that I have found
in the disseminated belt would unquestionably be rated as argentiferous
by a Western miner, but the assay showed that the structure in this
case was due to other causes, as only about two ounces were found.
An apparent exception was reported at the Peach Orchard diggings,
in Washington county, in the higher or Potosi member of the third
magnesian limestone, where Arthur Thacher found sulphide and carbonate
ore carrying 8 to 10 oz. of silver per ton; and a short-lived hamlet,
known as Silver City, sprang up to work them. I found, however, that
these deposits are associated with little vertical fissure-veins or
seams that unquestionably come up from the underlying porphyry.
Recently I examined the Jackson Revel mine, which has been considered
a silver mine for the last fifty years. It lies about seven miles
south of Fredericktown, and is a fissure-vein in Algonkian felsite,
where it protrudes, as a low hill, through the disseminated limestone
formation. A shaft has just been sunk about 150 ft. at less than
1000 ft. from the feather edge of the limestone. The vein is narrow,
only one to twelve inches wide, with slicken-sided walls, runs about
N. 20 deg. E., and dips 80 to 86 deg. eastward. White quartz forms
the principal part of the filling; the vein contains more or less
galena and zinc blende. Assays of the clean galena made by Prof. W. B.
Potter show only 2.5 oz. silver per ton, or no more than is frequently
found in the disseminated lead ores. As the lead in this fissure vein
may be regarded as of undoubted deep origin, and it is practically
non-argentiferous, this would seem to remove the last objection to
the theory of the deep-seated source of the lead in the disseminated
deposits of southeast Missouri.
MINING IN SOUTHEASTERN MISSOURI
BY WALTER RENTON INGALLS
(February 18, 1904)
The St. Joseph Lead Company, in the operation of its mines at Bonne
Terre, does not permit the cages employed for hoisting purposes to be
used for access to the mine. Men going to and from their work must
climb the ladders. This rule does not obtain in the other mines of the
district. The St. Joseph Lead Company employs electric haulage for
the transport of ore underground at Bonne Terre. In the other mines
of the district, mules are generally used. The flow of water in the
mines of the district is extremely variable; some have very little;
others have a good deal. The Central mine is one of the wettest in the
entire district, making about 2000 gal. of water per minute. Coal in
southeastern Missouri costs $2 to $2.25 per ton delivered at the mines,
and the cost of raising 2000 gal. of water per minute from a depth
of something like 350 ft. is a very considerable item in the cost of
mining and milling, which, in the aggregate, is expected to come to not
much over $1.25 per ton.
The ore shoots in the district are unusually large. Their precise trend
has not been identified. Some consider the predominance of trend to be
northeast; others, northwest. They go both ways, and appear to make
the greatest depositions of ore at their intersections. However, the
network of shoots, if that be the actual occurrence, is laid out on a
very grand scale. Vertically there is also a difference. Some shafts
penetrate only one stratum of ore; others, two or three. The orebody
may be only a few feet in thickness; it may be 100 ft. or more. The
occurrence of several overlying orebodies obviously indicates the
mineralization of different strata of limestone, while in the very
thick orebodies the whole zone has apparently been mineralized.
The grade of the ore is extremely variable. It may be only 1 or 2 per
cent. mineral, or it may be 15 per cent. or more. However, the average
yield for the district, in large mines which mill 500 to 1200 tons of
ore per day, is probably about 5 per cent. of mineral, assaying 65 per
cent. Pb, which would correspond to a yield of 3.25 per cent. metallic
lead in the form of concentrate. The actual recovery in the dressing
works is probably about 75 per cent., which would indicate a tenure of
about 4.33 per cent. lead in the crude ore.
LEAD MINING IN SOUTHEASTERN MISSOURI
BY R. D. O. JOHNSON
(September 16, 1905)
The lead deposits of southeastern Missouri carry galena disseminated
in certain strata of magnesian limestone. Their greater dimensions
are generally horizontal, but with outlines extremely irregular. The
large orebodies consist usually of a series of smaller bodies disposed
parallel to one another. These smaller members may coalesce, but are
generally separated from one another by a varying thickness of lean ore
or barren rock. The vertical and lateral dimensions of an orebody may
be determined with a fair degree of accuracy by diamond drilling, and a
map may be constructed from the information so obtained. Such a map (on
which are plotted the surface contours) makes it possible to determine
closely the proper location of the shaft, or shafts, considering also
the surface and underground drainage and tramming.
The first shafts in the district were sunk at Bonne Terre, where the
deposits lie comparatively near the surface. The early practice at this
point was to sink a number of small one-compartment shafts. As the
deposits were followed deeper, this gave way to the practice of putting
down two-compartment shafts equipped much more completely than were the
shallower shafts.
At Flat River (where the deposits lie at much greater depths, some
being over 500 ft.) the shafts are 7 × 14 ft., 6½ × 18 ft., and 7 × 20
ft. These larger dimensions give room not only for two cage-ways and a
ladder-way, but also for a roomy pipe-compartment. The large quantities
of water to be pumped in this part of the district make the care of
the pipes in the shafts a matter of first importance. At Bonne Terre
only such a quantity of water was encountered as could be handled by
bailing or be taken out with the rock; there the only pipe necessary
was a small air-pipe down one corner of the shaft. When the quantity
of water encountered is so great that the continued working of the mine
depends upon its uninterrupted removal, the care of the pipes becomes a
matter of great importance. A shaft which yields from 4000 to 5000 gal.
of water per minute is equipped with two 12 in. column pipes and two 4
in. steam pipes covered and sheathed. Moreover, the pipe compartment
will probably contain an 8 in. air-pipe, besides speaking-tubes, pipes
for carrying electric wires, and pipes for conducting water from upper
levels to the sump. To care for these properly there are required a
separate compartment and plenty of room.
Shafts are sunk by using temporary head frames and iron buckets of from
8 to 14 cu. ft. capacity. Where the influx of water was small, 104 ft.
have been sunk in 30 days, with three 8 hour shifts, two drills, and
two men to each drill; 2¾ in. drills are used almost exclusively; 3¼
in. drills have been used in sinking, but without apparent increase in
speed.
The influence of the quantity of water encountered upon the speed of
sinking (and the consequent cost per foot) is so great that figures are
of little value. Conditions are not at all uniform.
At some point (usually before 200 ft. is reached) a horizontal opening
will be encountered. This opening invariably yields water, the amount
following closely the surface precipitation. It is the practice to
establish at this point a pumping station. The shaft is “ringed” and
the water is directed into a sump on the side, from which it is pumped
out. This sump receives also the discharge of the sinking pumps.
The shafts sunk in solid limestone require no timbering other than that
necessary to support the guides, pipes, and ladder platforms. These
timbers are 8 × 8 in. and 6 x 8 in., spaced 7 or 8 ft. apart.
Shafts are sunk to a depth of 10 ft. below the point determined upon
as the lower cage landing. From the end at the bottom a narrow drift
is driven horizontally to a distance of 15 ft.; at that point it is
widened out to 10 ft. and driven 20 ft. further. The whole area (10 ×
20 ft.) is then raised to a point 28 or 30 ft. above the bottom of the
drift from the shaft. The lower part of this chamber constitutes the
sump. Starting from this chamber (on one side and at a point 10 ft.
above the cage landing, or 20 ft. above the bottom of the sump), the
“pump-house” is cut out. This pump-house is cut 40 ft. long and is as
wide as the sump is long, namely, 20 ft. A narrow drift is driven to
connect the top of the pump-house with the shaft. Through this drift
the various pipes enter the pump-house from the shaft.
The pumps are thus placed at an elevation of 10 ft. above the bottom
of the mine. Flooding of mines, due to failure of pumps or to striking
underground bodies of water, taught the necessity of placing the pumps
at such an elevation that they would be the last to be covered, thus
giving time for getting or keeping them in operation. The pumps are
placed on the solid rock, the air pumps and condensers at a lower level
on timbers over the sump.
With this arrangement, the bottom of the shaft serves as an antechamber
for the sump, in which is collected the washing from the mine and the
dripping from the shaft. The sump proper rarely needs cleaning.
The pumps are generally of high-grade, compound-and triple-expansion,
pot-valved, outside-packed plunger pattern. Plants with electrical
power distribution have recently installed direct-connected compound
centrifugal pumps with entire success.
Pumps of the Cornish pattern have never been used much in this region.
One such pump has been installed, but the example has not been followed
even by the company putting it in.
The irregular disposition of the ore renders any systematic plan of
drifting or mining (as in coal or vein mining) impossible. On each
side of the shaft and in a direction at right angles to its greater
horizontal dimension, drifts 12 to 14 ft. in width are driven to a
distance of 60 or 70 ft. In these broad drifts are located the tracks
and also the “crossovers” for running the cars on and off the cage.
When a deposit is first opened up, it is usually worked on two, and
sometimes three, levels. These eventually cut into one another, when
the ore is hoisted from the lower level alone.
The determination of the depth of the lower level is a matter of
compromise. Much good ore may be known to exist below; when it comes
to mining, it will have to be taken out at greater expense; but the
level is aimed to cut through the lower portions of the main body. It
is generally safe to predict that the ore lying below the upper levels
will eventually be mined from a lower level without the expense of
local underground hoisting and pumping.
The ore has simply to be followed; no one can say in advance how it
is going to turn out. The irregularity of the deposits renders any
general plan of mining of little or no value. Some managers endeavor to
outline the deposits by working on the outskirts, leaving the interior
as “ore reserves.” Such reserves have been found to be no reserves at
all, though the quality of the rock may be fairly well determined by
underground diamond drilling. Many of the deposits are too narrow to
permit the employment of any system of outlining and at the same time
keeping up the ore supply.
The individual bodies constituting the general orebody are rarely,
if ever, completely separated by barren rock; some “stringers” or
“leaders” of ore connect them. The careful superintendent keeps a
record on the monthly mine map of all such occurrences, or otherwise,
or of blank walls of barren rock that mark the edge of the deposit.
This precaution finds abundant reward when the drills commence to “cut
poor,” and when a search for ore is necessary.
The method of mining is to rise to the top of the ore and to carry
forward a 6 ft. breast. If the ore is thick enough, this is followed by
the underhand stope. Drill holes in the breast are usually 7 or 8 ft.
in depth; stope holes, 10 to 14 feet.
Both the roof and the floor are drilled with 8 or 10 ft. holes placed
8 or 10 ft. apart. These serve to prospect the rock in the immediate
neighborhood; in the roof they serve the further very important purpose
of draining out water that might otherwise accumulate between the
strata and that might force them to fall. The condition or safety
of the roof is determined by striking with a hammer. If the sound
is hollow or “drummy,” the roof is unsafe. If water is allowed to
accumulate between the loose strata, obviously it is not possible to
determine the condition of the roof.
It is the duty of two men on each shift to keep the mine in a safe
condition by taking down all loose and dangerous masses of rock. These
men are known as “miners.” It sometimes happens that a considerable
area of the roof gets into such a dangerous condition that it is either
too risky or too expensive to put in order, in which case the space
underneath is fenced off. As a general thing, the mines are safe and
are kept so. There are but few accidents of a serious nature due to
falling rock.
The roof is supported entirely by pillars; no timbering whatever is
used. The pillars are parts of the orebody or rock that is left. They
are of all varieties of size and shape. They are usually circular in
cross-section, 10 to 15 ft. in diameter and spaced 20 to 35 ft. apart,
depending upon the character of the roof. Pillars generally flare at
the top to give as much support to the roof as possible. The hight of
the pillars corresponds, of course, to the thickness of the orebody.
All drilling is done by 2¾ in. percussion drills. In the early days,
when diamonds were worth $6 per carat, underground diamond drills were
used. Diamond drills are used now occasionally for putting in long
horizontal holes for shooting down “drummy” roof. Air pressure varies
from 60 to 80 lb. Pressures of 100 lb. and more have been used, but the
repairs on the drills became so great that the advantages of the higher
pressure were neutralized.
Each drill is operated by two men, designated as “drillers,” or “front
hand” and “back hand.” The average amount of drilling per shift of 10
hours is in the neighborhood of 40 ft., though at one mine an average
of 55 ft. was maintained.
In some of the mines the “drillers” and “back hands” do the loading and
firing; in others that is done by “firers,” who do the blasting between
shifts. When the drillers do the firing, there is employed a “powder
monkey,” who makes up the “niphters,” or sticks of powder, in which are
inserted and fastened the caps and fuse; 35 per cent. powder is used
for general mining.
Battery firing is employed only in shaft sinking. In the mining work
this is found to be much more expensive; the heavy concussions loosen
the stratum of the roof and make it dangerous.
Large quantities of oil are used for lubrication and illumination.
“Zero” black oil and oils of that grade are used on the drills. Miners’
oil is generally used for illumination, though some of the mines use a
low grade of felsite wax.
Two oil cans (each holding about 1½ pints) are given to each pair of
drillers, one can for black oil and one for miners’ oil. These cans,
properly filled, are given out to the men, as they go on shift, at the
“oil-house,” located near the shaft underground. This “oil-house” is in
charge of the “oil boy,” whose duty it is to keep the cans clean, to
fill them and to give them out at the beginning of the shift. The cans
are returned to the oil-house at the end of the shift.
Kerosene is used in the hat-lamps in wet places.
The “oil-houses” are provided with three tanks. In some instances these
tanks are charged through pipes coming down the shaft from the surface
oil-house. These tanks are provided with oil-pumps and graduated
gage-glasses.
Shovelers or loaders operate in gangs of 8 to 12, and are supervised by
a “straw boss,” who is provided with a gallon can for illuminating oil.
The cars are 20 cu. ft. (1 ton) capacity. Under ordinary conditions one
shoveler will load 20 of these cars in a shift of 10 hours. They use
“half-spring,” long-handled, round-pointed shovels.
Cars are of the solid-box pattern, and are dumped in cradles. Loose
and “Anaconda” manganese-steel wheels are the most common. Gage of
track, 24 to 30 in., 16 lb. rails on main lines and 12 lb. on the side
and temporary tracks. Cars are drawn by mules. One mine has installed
compressed-air locomotives and is operating them with success.
Shafts are generally equipped with geared hoists, both steam and
electrically driven. Later hoists are all of the first-motion pattern.
Generally the cars are hoisted to the top and dumped with cradles. One
shaft, however, is provided with a 5-ton skip, charged at the bottom
from a bin, into which the underground cars are dumped. Upon arriving
at the top the skip dumps automatically. This design exhibits a number
of advantages over the older method and will probably find favor with
other mine operators.
THE LEAD ORES OF SOUTHWESTERN MISSOURI
BY C. V. PETRAEUS AND W. GEO. WARING
(October 21, 1905)
The lead ore of southwestern Missouri, and the adjoining area in the
vicinity of Galena, Kan., is obtained as a by-product of zinc mining,
the galena being separated from the blende in the jigging process.
Formerly the galena (together with “dry-bone,” including cerussite and
anglesite) was the principal ore mined from surface deposits in clay,
the blende being the subsidiary product. In the deeper workings blende
was found largely to predominate; this is shown by the shipments of the
district in 1904, which amounted to 267,297 tons of zinc concentrate
and 34,533 tons of lead concentrate.
The lead occurs in segregated cubes, from less than one millimeter up
to one foot in diameter. The cleavage is perfect, so that each piece
of ore when struck with a hammer breaks up into smaller perfect cubes.
In this respect the ore differs from the galena encountered in the
Rocky Mountain regions, where torsional or shearing strains seem in
most instances to have destroyed the perfect cleavage of the minerals
subsequent to their original deposition. Cases of schistose and twisted
structure occur in lead deposits of the Joplin district but rarely, and
they are always quite local.
The separation of the galena from the blende and marcasite (“mundic”)
in the ordinary process of jigging is very complete; the percentage
of zinc and iron in the lead concentrate is insignificant. As an
illustration of this, the assays of 100 recent consecutive shipments of
lead ore from the district, taken at random, are cited as follows:
7 shipments assayed from 57 to 70% lead
15 shipments assayed from 70 to 75% lead
46 shipments assayed from 75 to 79% lead
32 shipments assayed from 80 to 84.4% lead
Average of 100 shipments 78.4% lead
Fourteen shipment samples, ranging from 70 to 84.4 per cent. lead, were
tested for zinc and iron. These averaged 2.24 per cent. Fe and 1.78
per cent. Zn, the highest zinc content being 4.5 per cent. No bismuth
or arsenic, and only very minute traces of antimony, have ever been
found in these ores. They contain only about 0.0005 per cent. of silver
(one-seventh of an ounce per ton) and scarcely more than that of copper
(occurring as chalcopyrite).
The pig lead produced from these ores is therefore very pure, soft and
uniform in quality, so that the term “soft Missouri lead” has become a
synonym for excellence in the manufacture of lead alloys and products,
such as litharge, red and white lead, and orange mineral. Its freedom
from bismuth, which is generally present in Colorado lead, makes it
particularly suitable for white lead; also for glass-maker’s litharge
and red lead. These oxides, for use in making crystal glass, must be
made by double refining so as to remove even the small quantities
of silver and copper that are present. The resulting product, made
from soft Missouri lead, is far superior to any refined lead produced
anywhere in this country or in Europe, even excelling the famous
Tarnowitz lead. It gives a luster and clarity to the glass that no
other lead will produce. Lead from southeastern Missouri, Kentucky,
Illinois, Iowa, and Wisconsin yields identical results, but the
refining is more difficult, not only because the lead contains a little
more silver and copper, but also because it contains more antimony.
The valuation of the lead concentrate produced in the Joplin district
is based upon a wet assay, usually the molybdate or ferrocyanide
method. The price paid is determined variously. One buyer pays a
fixed price for average ore, making no deductions; as, for example,
at present rates, $32.25 per 1000 lb. whether the ore assays 75 or 84
per cent. Pb, pig lead being worth $4.75 at St. Louis.[6] Another pays
$32.25 for 80 per cent. ore, or over, deducting 50c. per unit for ores
assaying under 80 per cent. Another pays for 90 per cent. of the lead
content of the ore as shown by the assay, at the St. Louis price of pig
lead, less a smelting charge of, say, $6 to $8 per ton of ore.
The history of the development of lead ore buying in the Joplin
district is rather curious. In the early days of the district the ore
was smelted wholly on Scotch hearths, which, with the purest ores,
would yield 70 per cent. metallic lead. No account was taken of the
lead in the rich slag, chemical determinations being something unknown
in the district at that time; it being supposed generally that pure
galena contained 700 lb. lead to the 1000 lb. of ore, the value of
700 lb. lead, less $4.50 per 1000 lb. of ore for freight and smelting
costs, was returned to the miner. The buyers graded the ore, according
to their judgment, by its appearance, as to its purity and also as to
its behavior in smelting; an ore, for example, from near the surface,
imbedded in the clay and coated more or less with sulphate, yielded its
metal more freely than the purer galenas from deeper workings.
This was the origin of the present method of buying—a system that would
hardly be tolerated except for the fact that the lead is, as previously
stated, considered a by-product of zinc mining.
Originally all the lead ore from the Missouri-Kansas district was
smelted in the same region, either in the air furnace (reverberatory
sweating-furnace) or in the water-back Scotch hearth. Competition
gradually developed in the market. Lead refiners found the pure
sulphide of special value in the production of oxidized products.
Some of the ore found its way to St. Louis, and even as far away as
Colorado, where it was used to collect silver. Since the formation of
the American Smelting and Refining Company and the greatly increased
output of the immense deposits of lead ore in Idaho, no Missouri lead
ore has gone to Colorado.
Up to 1901, one concern had more or less the control of the
southwestern Missouri ores. At the present time, lead ore is bought
for smelters in Joplin, Carterville, and Granby, Mo., Galena, Kan.,
and Collinsville, Ill., and complaint is heard that present prices are
really too high for the comfort of the smelters. Yet the old principle
of paying for lead ores upon the supposed yield of 70 per cent.,
irrespective of the real lead content, is still largely in vogue.
Any one interested in the matter will find it quite instructive to
calculate the returning charges, or gross profits, in smelting these
ores, on the basis of 70 per cent. recovery, at $32.25 per 1000 lb. of
ore, less 50c. per ton haulage, with lead at $4.77 per 100 lb. at St.
Louis. No deduction, it should be remarked, is ever made for moisture
in lead ores in this district. It is of interest to observe that
Dr. Isaac A. Hourwich estimates (in the U. S. Census Special Report
on Mines and Quarries recently issued) the average lead contents of
the soft lead ores of Missouri in 1902 at 68.2 per cent., taking as a
basis the returns from five leading mining and smelting companies of
Missouri, which reported a product of 70,491 tons of lead from 103,428
tons of their own and purchased ore. The average prices for lead ore
in 1902 were reported as follows, per 1000 lb.: Illinois, $19.53;
Iowa, $24.48; Kansas, $23.51; Missouri, $22.17; Wisconsin, $23.29;
Rocky Mountain and Atlantic Coast States, $10.90. In 1903, according
to Ingalls (“The Mineral Industry,” Vol. XII), the ore from the Joplin
district commanded an average price of $53 per 2000 lb., while the
average in the southeastern district was $46.81.
PART II
ROAST-REACTION SMELTING
SCOTCH HEARTHS AND REVERBERATORY FURNACES
LEAD SMELTING IN THE SCOTCH HEARTH
BY KENNETH W. M. MIDDLETON
(July 6, 1905)
In view of the fact that the Scotch hearth in its improved form is now
coming to the front again to some extent in lead smelting, it may prove
interesting to give a brief account of its present use in the north of
England.
Admitting that, where preliminary roasting is necessary, the best
results can be obtained with the water-jacketed blast furnace (this
being more especially the case where labor is an expensive item), we
have still as an alternative the method of smelting raw in the Scotch
hearth. At one works, which I recently visited, all the ore was smelted
raw; at another, all the ore received a preliminary roast, and it is
instructive to compare the results obtained in the two cases. The
following data refer to a fairly “free-smelting” galena assaying nearly
80 per cent. of lead.
When smelting raw ore in the hearth, fully 7½ long tons can be treated
in 24 hours, the amount of lead produced direct from the furnace in the
first fire being 8400 to 9000 lb.; this is equivalent to 56 to 60 per
cent. of lead, the remaining 24 to 20 per cent. going into the fume and
the slag.
When smelting ore which has received a preliminary roast of two hours,
12,000 lb. of lead is produced direct from the hearth, this being
equivalent to 65 per cent. of the ore. When the ore is roasted, the
output of the hearth is practically the same for all ores of equal
richness; but when smelting raw, if the galena is finely divided, the
output may fall much below that given herewith; while, on the other
hand, under the most favorable conditions it may rise to 12,000 lb. in
24 hours, or even more.
I had an opportunity of seeing a parcel of galena carrying 84 per cent.
of lead (but broken down very fine) smelted raw. The ore was kept damp
and the blast fairly low; but, in spite of that, a quantity of the ore
was blown into the flue, and only 5100 lb. of lead was produced from
the hearth in 24 hours.
Galena carrying only 65 per cent. of lead does not give nearly as
satisfactory results when smelted raw in the hearth; barely six tons of
ore can be smelted in 24 hours, and only 4500 to 5400 lb. of lead can
be produced directly. This is equivalent to, say, 43 per cent. of the
ore in the first fire; the remaining 22 per cent. goes into the slag or
to the flue as fume. Moreover, the 65 per cent. ore requires 1500 lb.
of coal in 24 hours, while the 80 per cent. galena uses only 1000 lb.
Turning now for a moment to the costs of smelting raw and of smelting
after a preliminary roast, we find that (in the case of the two works
we have been considering) the results are all in favor of smelting raw,
so far as a galena carrying nearly 80 per cent. is concerned.
The cost of smelting, per ton of lead produced, is given herewith:
ORE SMELTED RAW
Smelters’ wages $2.04
“ coal (425 lb.) 0.38
——-
Total $2.42
A very small quantity of lime is also used in this case for some ores,
but its cost would never amount to more than 4c. per ton of lead
produced.
ORE RECEIVING A PRELIMINARY ROAST
Roasters’ wages $0.61
“ coal (425 lb.) 0.65
Smelters’ wages 1.08
“ coal (75 lb.) 0.11
Peat and lime 0.08
——-
Total $2.53
It should be noted also that the smelters at the works where the ore
was not roasted receive higher pay. In the eight-hour shift they
produce about 1½ tons of lead; and as there are two of them to a
furnace, they make $3.06 between them, or $1.53 each. The two men
smelting roasted ore produce about two tons in an eight-hour shift, and
therefore each receives $1.08 per shift.
Coming now to fume-smelting in the hearth, we can again compare the
results obtained in smelting raw and after roasting. It is well to
bear in mind, also, that, while only 6½ per cent. of the lead goes in
the fume when smelting roasted ores in the hearth, a considerably
larger proportion is thus lost when smelting raw ores. When fume is
smelted raw, it is best dealt with when containing about 40 per cent.
of moisture. One man attends to the hearth (instead of two as when
smelting ore), and in 24 hours 3000 lb. of lead is produced, the amount
of coal used being 2100 lb. No lime is required.
When smelting roasted fume, two men attend to the hearth and the output
is 6000 lb. in 24 hours, the amount of coal used being 1800 lb. In this
latter case fluorspar happens to be available (practically free of
cost), and a little of it is used with advantage in fume-smelting, as
well as a small quantity of lime.
The cost of fume-smelting per ton of lead produced is given herewith:
FUME SMELTED RAW
Smelters’ wages $2.88
“ coal (1400 lb.) 2.13
——-—-
Total $5.01
FUME RECEIVING A PRELIMINARY ROAST
Roasters’ wages $2.08
“ coal (1450 lb.) 2.18
Smelters’ wages 2.04
“ coal (600 lb.) 0.92
Peat and lime 0.08
———--
Total $7.30
In this case, as in that of ore, the smelter of the raw fume gets
better pay; he has $1.44 per eight-hour shift, while the smelter of the
roasted ore has only $1.02 per eight-hour shift.
Fume takes four hours to roast, as compared to the two hours taken by
ore.
From these facts regarding Scotch-hearth smelting, it would seem that
with galena carrying, say, over 70 per cent. lead (but more especially
with ore up to 80 per cent. in lead, and, moreover, fairly free from
impurities detrimental to “free” smelting), very satisfactory results
can be obtained by smelting raw. Against this, however, it must be said
that at the works where the ore is roasted attempts at smelting raw
have been made several times without sufficient success to justify the
adoption of this method, although the ores smelted average 75 per cent.
lead and seem quite suitable for the purpose.
Probably this may be accounted for by the fact that the method of
running the furnace when raw ore is being smelted is rather different
from that adopted when dealing with roasted ore. Moreover, at the works
under notice the furnaces are not of the most modern construction; and,
as the old custom of dropping a peat in front of the blast every time
the fire is made up still survives, it is necessary to shut off the
blast while this is being done, and the fire is then apt to get rather
slack.
The gray slag produced in the hearth is smelted in a small blast
furnace, a little poor fume, and sometimes a small quantity of
fluorspar, being added to facilitate the process. Some figures
regarding slag-smelting may be of interest. The slag-smelters produce
9000 lb. of lead in 24 hours. The cost of slag-smelting per ton of lead
produced is as follows:
Smelters’ wages $1.60
Coke (1500 lb.) 3.42
Peat 0.06
———--
Total $5.08
Recent analyses of Weardale (Durham county) lead smelted in the Scotch
hearth, and slag-lead smelted in the blast furnace, are given herewith:
─────────┬───────────────────┬────────────────────┬──────────────────
│ FUME-LEAD FROM │ SILVER-LEAD FROM │ SLAG-LEAD FROM
│ HEARTH │ HEARTH │ BLAST FURNACE
─────────┼───────────────────┼────────────────────┼──────────────────
Lead │ 99.957 │ 99.957 │ 99.013
Silver │ 0.0035 │ 0.0200 │ 0.0142
│ (1oz. 2dwt. 21gr. │ (6oz. 10dwt. 16gr. │(4oz. 12dwt. 18gr.
│ per Long Ton) │ per Long Ton) │ per Long Ton)
Tin │ nil │ nil │ nil
Antimony │ nil │ nil │ 0.874
Copper │ nil │ nil │ 0.024
Iron │ 0.019 │ 0.019 │ 0.023
Zinc │ nil │ nil │ nil
│ ──────── │ ──────── │ ────────
│ 99.9795 │ 99.9960 │ 99.9482
─────────┴───────────────────┴────────────────────┴──────────────────
The ordinary form of the Scotch hearth is probably too well known
to need much description. The dimensions which have been found most
suitable are as follows: Front to back, 21 in.; width, 27 in.; depth
of hearth, 8 to 12 in. Formerly the distance from front to back was 24
in., but this was found too much for the blast and for the men.
The cast-iron hearth which holds the molten lead is set in brickwork;
if 8 in. deep and capable of holding about ¾ ton of lead, it is quite
large enough. The workstone or inclined plate in front of the hearth
is cast in one piece with it, and has a raised holder on either side
at the lower edge, and a gutter to convey the overflowing lead to the
melting-pot. The latter is best made with a partition and an opening
at the bottom through which clean lead can run, so that it can be
ladled into molds without the necessity for skimming the dross off
the surface. It is well also to have a small fireplace below the
melting-pot.
On each side of the hearth, and resting on it, is a heavy cast-iron
block, 9 in. thick, 15 in. high, 27 to 28 in. long. To save metal,
these are now cast hollow and air is caused to pass through them. On
the back of the hearth stands another cast-iron block known as the
“pipestone,” through which the blast comes into the furnace. In the
older forms of pipestone the blast comes in through a simple round or
oval pipe, a common size being 3 or 4 in. wide by 2½ in. high, and the
pipestone is not water-cooled. With this construction the hearth will
not run satisfactorily unless the pipestone is set with the greatest
care, so as to have the tuyere exactly in the center, and as there
is no water-cooling the metal quickly burns away when fume is being
smelted. Moreover, the blast is apt to be stopped by slag adhering to
the end of the pipe. As already mentioned, a peat is dropped in front
of the blast every time the fire is made up, with the object of keeping
a clear passage open for the blast. This old custom has, however,
several serious disadvantages; first, it prevents the blast being kept
on continuously; and, second, it makes it necessary to have the hearth
open at the top so that the smelter-man can go in by the side of it. In
this case the ore is fed from the side by the smelter-man, who works
under the large hood placed above the furnace to carry away the fume.
Even when he is engaged in shoveling back the fire from the front and
is not underneath the hood, it is impossible to prevent some fume from
blowing out; and there is much more liability to lead-poisoning than
when the hearth is closed at the top by the chimney and the smelter-men
work from the front. The best arrangement is to have the hearth
entirely closed in by the chimney, except for the opening at the front,
and to have a small auxiliary flue above the workstone leading direct
to the open air to catch any fume that may blow out past the shutter in
front of the hearth.
In an improved form of pipestone, a pipe connected to the blast-main
fits into the semicircular opening at the back and is driven tight
against a ridge in the flat side of the opening. Going through the
pipestone, the arch becomes gradually flatter, and the blast emerges
into the hearth, about 2 in. above the level of the molten lead,
through an oblong slit 12 in. long by 1 in. wide, with a ledge
projecting 1½ in. immediately above it. The back and front are similar,
so that when one side gets damaged the pipestone can be turned back to
front.
Water is conveyed in a 2½ in. iron pipe to the pipestone, and after
passing through it is led away from the other end to a water-box, which
stands beside the hearth and into which the red-hot lumps of slag are
thrown to safeguard the smelters from the noxious fumes.
On the top of the pipestone rests an upper backstone, also of cast
iron; it extends somewhat higher than the blocks at the sides. All this
metal above the level of the lead is necessary because the partially
fused lumps which stick to it have to be knocked off with a long bar,
so that if fire-bricks were used in place of cast iron they would soon
be broken up and destroyed.
With a covered-in hearth, when the ore is charged from the front,
the following is the method adopted in smelting raw ore: The charge
floats on the molten lead in the hearth, and at short intervals the
two smelters running the furnace ease it up with long bars, which they
insert underneath in the lead. Any pieces of slag adhering to the sides
and pipestone are broken off. After easing up the fire, the lumps of
partially reduced ore, mixed with cinders and slag, are shoveled on
to the back of the fire; the slag is drawn out upon the workstone
(any pieces of ore adhering to it being broken off and returned to
the hearth), and it is then quenched in a water-box placed alongside
the workstone. One or two shovelfuls of coal, broken fairly small
and generally kept damp, are thrown on the fire, together with the
necessary amount of ore, which is also kept damp if in a fine state
of division. It is part of the duty of the two smelters to ladle out
the lead from the melting-pot into the molds. In smelting ore a fairly
strong, steady blast is required, and it is made to blow right through
so as to keep the front of the fire bright. A little lime is thrown on
the front of the fire when the slag gets too greasy.
When smelting raw fume one man attends to the furnace. It does not
have to be made up nearly as frequently, the work being easier for
one man than smelting ore is for two. The unreduced clinkers and slag
are dealt with exactly as in smelting ore; and coal is also, in this
case, thrown on the back of the fire, but the blast does not blow
right through to the front. On the contrary, the front of the fire is
kept tamped up with fume, which should be of the coherency of a thick
mud. The blast is not so strong as that necessary for ore. The idea is
partially to bake the fume before submitting it to the hottest part of
the furnace, or to the part where the blast is most strongly felt. It
is only when smelting fume that it is necessary to keep the pipestone
water-cooled.
To start a furnace takes from two to three hours. The hearth is left
full of lead, and this has to be melted before the hearth is in normal
working order. Drawing the fire takes about three-quarters of an hour;
the clinkers are taken off and kept for starting the next run, and the
sides and back of the hearth are cleaned down.
THE FEDERAL SMELTING WORKS, NEAR ALTON, ILL.[7]
BY O. PUFAHL
(June 2, 1906)
The works of the Federal Lead Company, near Alton, Ill., were erected
in 1902. They have a connection with the Chicago, Peoria & St. Louis
Railway, by which they receive all their raw materials, and by which
all the lead produced is shipped.
The ore smelted is galena, with dolomitic gangue, and a small quantity
of pyrites (containing a little copper, nickel, and cobalt) from
southeastern Missouri, and consists chiefly of fine concentrates,
containing 60 to 70 per cent. lead. In addition thereto a small
proportion of lump ore is also smelted.
A striking feature at these works is the excellent facility for
handling the materials. The bins for the ore, coke and coal are made
of concrete and steel and are filled from cars running on tracks
laid above them. For transporting the materials about the works a
narrow-gage railway with electric locomotives is used.
The ores are smelted by the Scotch-hearth process. There are 20 hearths
arranged in a row in a building constructed wholly of steel and stone.
The sump (4 × 2 × 1 ft.) of each furnace contains about one ton of
lead. The furnaces are operated with low-pressure blast from a main
which passes along the whole row. The blast enters the furnace from a
wind chest at the back through eight 1 in. iron pipes, 2 in. above the
bath of lead. The two sides and the rear wall are cooled by a cast-iron
water jacket of 1 in. internal width.
Two men work, in eight-hour shifts, at each of the furnaces, receiving
4.75 and 4.25c. respectively for every 100 lb. of lead produced. The
ore is weighed out and heaped up in front of the furnaces; on the
track near by the coke is wheeled up in a flat iron car with two
compartments. The furnacemen are chiefly negroes. At the side of each
furnace is a small stock of coal, which is used chiefly for maintaining
a small fire under the lead kettle. Only small quantities of coal are
added from time to time during the smelting operation.
Over each furnace is placed an iron hood, through which the fumes and
gases escape. They pass first through a collecting pipe, extending
through the whole works, to a 1500 ft. dust flue, measuring 10 × 10
ft., in internal cross-section. Near the middle of this is placed a
fan of 100,000 cu. ft. capacity per minute, which forces the fumes and
gases into the bag-house, where they are filtered through 1500 sacks of
loosely woven cotton cloth, each 25 ft. long and 18 in. in diameter,
and thence pass up a 150 ft. stack.
The dust recovered in the collecting flue is burnt, together with the
fume caught by the bags, the coal which it contains furnishing the
combustible. It burns smolderingly and frits together somewhat. The
product (chiefly lead sulphate) is then smelted in a shaft furnace,
together with the gray slag from the hearth furnaces. The total
extraction of lead is about 98 per cent., i.e., the combined process
of Scotch-hearth and blast-furnace smelting yields 98 per cent. of the
lead contained in the crude ore.
The direct yield of lead from the Scotch hearths is about 70 per cent.
They also produce gray slag, containing much lead, which amounts to
about 25 per cent. of the weight of the ore. About equal proportions
of lead pass into the slag and into the flue dust. When working to
the full capacity, with rich ore (80 per cent. lead and more) the 20
furnaces can produce about 200 tons of lead in 24 hours. The coke
consumption in the hearth furnaces amounts to only 8 per cent. of the
ore. The lead from these furnaces is refined for 30 minutes to one
hour by steam in a cast-iron kettle of 35 tons capacity, and is cast
into bars either alone or mixed with lead from the shaft furnace. The
“Federal Brand” carries nearly 99.9 per cent. lead, 0.05 to 0.1 per
cent. copper, and traces of nickel and cobalt.
The working up of the between products from the hearth-furnaces is
carried out as follows: Slag, burnt flue dust and roasted matte from
a previous run, together with a liberal proportion of iron slag (from
the iron works at Alton), are smelted in a 12-tuyere blast furnace
for work-lead and matte. The furnace is provided with a lead well at
the back. The matte and slag are tapped off together at the front and
flow through a number of slag pots for separation. The shells which
remain adhering to the walls of the pots on pouring out the slag are
returned to the furnace. All the waste slag (containing about 0.5 per
cent. lead) is dumped down a ravine belonging to the territory of the
smeltery.
The lead from the shaft furnace is liquated in a small reverberatory
furnace, of which the hearth consists of two inclined perforated
iron plates. The residue is returned to the shaft furnace, while the
liquated lead flows directly to the refining kettle, which is filled
in the course of four hours. Here it is steamed for about one hour and
is then cast into bars through a Steitz siphon, after skimming off the
oxide. The matte is crushed and roasted in a reverberatory furnace (60
ft. long).
The power plant comprises three Stirling boilers and two 250 h. p.
compound engines, of which one is for reserve; also one steam-driven
dynamo, coupled direct to the engine, furnishing the current for the
entire plant, for the electric locomotives, etc.
The coke is obtained from Pennsylvania and costs about $4 a ton, while
the coal comes from near-by collieries and costs $1 per ton.
In the well-equipped laboratory the lead in the ores and slags is
determined daily by Alexander’s (molybdate) method, while the silver
content of the lead (a little over 1 oz. per ton) is estimated only
once a month in an average sample. When the plant is in full operation
it gives employment to 150 men. Cases of lead-poisoning are said to
occur but rarely, and then only in a mild form.
LEAD SMELTING AT TARNOWITZ
(September 23, 1905)
The account of the introduction of the Huntington-Heberlein process at
Tarnowitz, Prussia, published elsewhere in this issue, is of peculiar
interest inasmuch as it tells of the complete displacement by the new
process of one of the old processes of lead smelting which had become
classic in the art. The roast-reaction process of lead smelting,
especially as carried out in reverberatory furnaces, has been for a
long time decadent, even in Europe. Tarnowitz was one of the places
where it survived most vigorously.
Outside of Europe, this process never found any generally extensive
application. It was tried in the Joplin district, and elsewhere in
Missouri, with Flintshire furnaces in the seventies. Later it was
employed with modified Flintshire and Tarnowitz furnaces at Desloge,
in the Flat River district of Missouri, where the plant is still in
operation, but on a reduced scale.
The roast-reaction process of smelting, as practised at Tarnowitz,
was characterized by a comparatively large charge, slow roasting and
low temperature, differing in these respects from the Carinthian and
Welsh processes. It was not aimed to extract the maximum proportion of
lead in the reverberatory furnace itself, the residue therefrom, which
inevitably is high in lead, being subsequently smelted in the blast
furnace. Ores too low in lead to be suitable for the reverberatory
smelting were sintered in ordinary furnaces and smelted in the blast
furnace together with the residue from the other process. In both of
these processes the loss of lead was comparatively high. One of the
most obvious advantages of the Huntington-Heberlein process is its
ability to reduce the loss of lead. The result in that respect at
Tarnowitz is clearly stated by Mr. Biernbaum, whose paper will surely
attract a good deal of attention.[8]
LEAD SMELTING IN REVERBERATORY FURNACES AT DESLOGE, MO.
BY WALTER RENTON INGALLS
(December 16, 1905)
The roast-reaction method of lead smelting in reverberatory furnaces
never found any general employment in the United States, although
in connection with the rude air-furnaces it was early introduced in
Missouri. The more elaborate Flintshire furnaces were tried at Granby,
in the Joplin district, but they were displaced there by Scotch
hearths. The most extensive installation of furnaces of the Flintshire
type was made at Desloge, in the Flat River district of southeastern
Missouri. This continued in full operation until 1903, when the major
portion of the plant was closed, it being found more economical to ship
the ore elsewhere for smelting. However, two furnaces have been kept
in use to work up surplus ore. As a matter of historic interest, it is
worth while to record the technical results at Desloge, which have not
previously been described in metallurgical literature.
The Desloge plant, which was situated close to the dressing works
connected with the mine, and was designed for the smelting of its
concentrate, comprised five furnaces. The furnaces were of various
constructions. The oldest of them was of the Flintshire type, and
had a hearth 10 ft. wide and 14 ft. long. The other furnaces were a
combination of the Flintshire and Tarnowitz types. They were built
originally like the newer furnaces at Tarnowitz, Upper Silesia, with a
rather large rectangular hearth and a lead sump placed at one side of
the hearth near the throat end; but good results were not obtained from
that construction, wherefore the furnaces were rearranged with the sump
at one side, but in the middle of the furnace, as in the Flintshire
form. The rectangular shape of the Tarnowitz hearth was, however,
retained. Furnaces thus modified had hearths 11 ft. wide and 16 ft.
long, except one which had a hearth 13 ft. wide.
The same quantity of ore was put through each of these furnaces, the
increase in hearth area being practically of no useful effect, because
of inability to attain the requisite temperature in all parts of the
larger hearths with the method of heating employed. The men objected
especially to a furnace with hearth 13 ft. wide, which it was found
difficult to keep in proper condition, and also difficult to handle
efficiently. Even the width of 11 ft. was considered too great, and
preference was expressed for a 10 ft. width. In this connection, it may
be noted that the old furnaces at Tarnowitz were 11 ft. 9 in. long and
10 ft. 10 in. wide, while the new furnaces were 16 ft. long and 8 ft.
10 in. wide (Hofman, “Metallurgy of Lead,” fifth edition, p. 112). All
of these dimensions were exceeded at Desloge.
The Flintshire furnaces at Desloge had three working doors per side;
the others had four, but only three per side were used, the doors
nearest the throat end being kept closed because of insufficient
temperature in that part of the furnace. The furnace with hearth 11
× 14 ft. had a grate area of 6.5 × 3 ft. = 19.5 sq. ft.; the 11 × 16
furnaces had grates 8 × 3 = 24 ft. sq. The ratios of grate to hearth
area were therefore approximately 1:8 and 1:7.3, respectively. (Compare
with ratio of 1:10 at Tarnowitz, and 1:6⅔ at Stiperstones.) The ash
pits were open from behind in the customary English fashion. The grate
bars were cast iron, 36 in. long. The bars were 1 in. thick at the top,
with ⅝ in. spaces between them. The open spaces were 32 in. long,
including the rib in the middle. The bars were 4 in. deep at the middle
and 2 in. at the ends. The distance from the surface of the grate bars
to the fire-door varied in the different furnaces. Some of those with
hearths 11 × 16 ft. and grates 8 × 3 ft. had the bars 6 in. below the
fire-door; in others the bars were almost on a level with the fire-door.
The furnaces were run with a comparatively thin bed of coal on the
grate, and combustion was very imperfect, the percentage of unburned
carbon in the ash being commonly high. This was unavoidable with the
method of firing employed and the inferior character of the coal
(southern Illinois). The excessive consumption of coal was due largely,
however, to the practice of raking out the entire bed of coal at the
beginning of the operation of “firing down” (beginning the reaction
period), when a fresh fire was built with cordwood and large lumps of
coal.
Each furnace had two flues at the throat, 16 × 18 in. in size, each
flue being provided with a separate damper. Each furnace had an
iron chimney approximately 55 ft. high, of which 13 ft. was a brick
pedestal (64 × 64 in.) and the remaining 42 ft. sheet steel, guyed. The
chimneys were 42 in. in diameter. The distance from the outside end
of the furnace to the chimney was approximately 6 ft., and there was
consequently but little opportunity for flue dust to collect in the
flue. About once a month, however, the chimney was opened at the base
and about two wheelbarrows (say 600 lb.) of flue dust, assaying about
50 per cent. lead, was recovered per furnace.
The furnace house was a frame building 45 ft. wide, with boarded sides
and a corrugated-iron pitch roof, supported by steel trusses. The
furnaces were set in this house, side by side, their longitudinal axes
being at right angles to the longitudinal axis of the building. The
distance from the outside of the fire-box end of the furnace to the
side of the building was 10 ft. The coal was unloaded from a railway
track alongside of the building and was wheeled to the furnace in
barrows. Some of the furnaces were placed 18 ft. apart; others 22 ft.
apart. The men much preferred the greater distance, which made their
work easier, an important consideration in this method of smelting.
The hight from the floor to the working door of the furnace was
approximately 36 in. The working doors were formed with cast-iron
frames, making openings 7 × 11 in. on the inside and 15 × 28 in. on
the outside. On the side of the furnace opposite the middle working
door was placed a cast-iron hemispherical pot, set partially below the
floor-line. This pot was 16 in. deep and 24 in. in diameter; the metal
was ¼ in. thick. The distance from the top of the pot to the line of
the working door was 31 in.; from the top of the pot to the bottom of
the tap-door was 7 in. The tap-door was 4 in. wide and 9 in. high,
opening through a cast-iron plate 1½ in. thick. Below the tap-door
and on a line with the upper rim of the pot was a tap-hole 3½ in. in
diameter. The frames of the working doors had lugs in front, against
which the buckstaves bore, to hold the frames in position. All other
parts of the sides of the furnace, including the fire-box, were cased
with ⅝ in. cast-iron plates, which were obviously too light, being
badly cracked.
The cost of a furnace when built in 1893 was approximately $1400,
not including the chimney; but with the increased cost of material
the present expense would probably be about $2000. Notwithstanding
the light construction of the furnaces, repairs were never a large
item. Once a month a furnace was idle about 24 hours while the throat
was being cleaned out, and every two months some repairing, such as
relining the fire-boxes, etc., was required. If repairs had to be made
on the inside of the furnace, two days would be lost while it was
cooling sufficiently for the men to enter. In refiring a furnace, from
8 to 12 hours was required to raise it to the proper temperature. Out
of the 365 days of the year, a furnace would lose from 20 to 25 days,
for cleaning the throat and making repairs to the fire-box, arch, etc.
When a furnace was run with two shifts the schedule of operation was as
follows:
Drop charge 4 a.m.
Begin work 7 a.m.
Begin firing down 11 a.m.
Begin first tapping 1 p.m.
Rake out slag 2.30 p.m.
Begin second tapping 3 p.m.
Drop charge 4 p.m.
Begin working 5.30 p.m.
Begin firing down 11 p.m.
Begin first tapping 1 a.m.
Rake out slag 2.30 a.m.
Begin second tapping 3 p.m.
With three shifts on a furnace, the schedule was as follows:
Drop charge 7 a.m.
Begin firing down 12 a.m.
Begin tapping 1 p.m.
Rake out slag 2 p.m.
Begin tapping 2.30 p.m.
Drop charge 3 p.m.
Begin firing down 8 p.m.
Begin tapping 9 p.m.
Rake out slag 10 p.m.
Begin tapping 10.30 p.m.
Drop charge 11.00 p.m.
Begin firing down 4 a.m.
Begin tapping 5 a.m.
Rake out slag 6 a.m.
Begin tapping 6.30 a.m.
The hearths were composed of about 8 in. of gray slag beaten down
solidly on a basin of brick, which rested on a filling of clay, rammed
solid. The hearth was patched if necessary after the drawing of each
charge.
The system of smelting was analogous to that which was practiced
in Wales rather than to the Silesian, the charges being worked off
quickly, and with the aim of making a high extraction of lead directly
and a gray slag of comparatively low content in lead. The average
furnace charge was 3500 lb. At the beginning of the reaction period
about 85 to 100 lb. of crushed fluorspar was thrown into the furnace
and mixed well with the charge. The furnace doors were then closed
tightly and the temperature raised, the grate having previously been
cleaned. At the first tapping about 1200 lb. of lead would be obtained.
A small quantity of chips and bark was thrown into the lead in the
kettle, which was then poled for a few minutes, skimmed, and ladled
into molds, the pigs weighing 80 lb. The skimmings and dross were
put back into the furnace. The pig lead was sold as “ordinary soft
Missouri.” The gray slag was raked out of the furnace, at the end of
the operation, into a barrow, by which it was wheeled to a pile outside
of the building. Shipments of the slag were made to other smelters from
time to time, 95 per cent. of its lead content being paid for when its
assay was over 40 per cent., and 90 per cent. when lower.
Each furnace was manned by one smelter ($1.75) and one helper ($1.55)
per shift, when two shifts per 24 hours were run. They had to get their
own coal, ore and flux, and wheel away their gray slag and ashes. In
winter, when three shifts were run, the men were paid only $1.65 and
$1.50 respectively. There was a foreman on the day shift, but none at
night. The total coal consumption was ordinarily about 0.8 to 0.9 per
ton of ore. Run-of-mine coal was used, which cost about $2 per ton
delivered. The coal was of inferior quality, and it was wastefully
burned, as previously referred to, wherefore the consumption was high
in comparison with the average at Tarnowitz, where it used to be about
0.5 per ton of ore.
The chief features of the practice at Desloge are compared with those
at Tarnowitz, Silesia and Holywell (Flintshire), and Stiperstones
(Shropshire), Wales, in the following table, the data for Silesia and
Wales being taken from Hofman’s “Metallurgy of Lead,” fifth edition,
pp. 112, 113.
──────────────────────┬─────────┬────────┬─────────┬─────────┬────────
DETAIL │HOLYWELL │ STIPER-│TARNOWITZ│TARNOWITZ│ DESLOGE
│ │ STONES │ │ │
──────────────────────┼─────────┼────────┼─────────┼─────────┼────────
Hearth length, ft. │ 12.00 │ 9.75 │ 11.75 │ 16.00 │ 16.00
Hearth width, ft. │ 9.50 │ 9.50 │ 10.83 │ 8.83 │ 11.00
Grate length, ft. │ 4.50 │ 4.50 │ 8.00 │ 8.00 │ 8.00
Grate width, ft. │ 2.50 │ 2.50 │ 1.67 │ 1.67 │ 3.00
Grate area: hearth │ │ │ │ │
area │ 1:8 │ 1:6⅔ │ 1:10 │ 1:10 │ 1:7⅓
Charges per 24 hr., │ 3 │ 3 │ 2 │ 2 │ 3
Ore smelted per │ │ │ │ │
24 hr., lb. │ 7,050 │ 7,050 │ 8,800 │ 16,500 │ 10,500
Assay of ore, % Pb │ 75-80 │ 77.5 │ 70-74 │ 70-74 │ 70
Gray slag, % of charge│ 12 │ │ 15 │ 30 │ 27
Gray slag, % Pb │ 55 │ │ 38.8 │ 56 │ 38
Men per 24 hr. │ 6 │ 4 │ 4 │ 6 │ 6
Coal used per ton ore │0.57-0.76│ 0.56 │ 0.46 │ 0.50 │ 0.90
──────────────────────┴─────────┴────────┴─────────┴─────────┴────────
The regular furnace charge at Desloge was 3500 lb. The working of three
charges per 24 hours gave a daily capacity of 10,500 lb. per furnace.
These figures refer to the wet weight of the concentrate, which was
smelted just as delivered from the mill. Its size was 9 mm. and finer.
Assuming its average moisture content to be 5 per cent., the daily
capacity per furnace was about 10,000 lb. (5 tons) of dry ore.
The metallurgical result is indicated by the figures for two months
of operation in 1900. The quantity of ore smelted was 1012 tons,
equivalent to approximately 962 tons dry weight. The pig lead produced
was 523.3 tons, or 54.4 per cent. of the weight of the ore. The gray
slag produced was 262.25 tons, or about 27 per cent. of the weight of
the ore. The assay of the ore was approximately 70 per cent. lead,
giving a content of 673.4 tons in the ore smelted. The gray slag
assayed approximately 38 per cent. lead, giving a content of 99.66
tons. Assuming that 90 per cent. of the lead in the gray slag be
recoverable in the subsequent smelting in the blast furnace, or 89.7
tons, the total extraction of lead in the process was 523.3 + 89.7 ÷
673.4 = 91 per cent. The metallurgical efficiency of the process was,
therefore, reasonably high, especially in view of the absence of dust
chambers.
* * * * *
The cost of smelting with five furnaces in operation, each treating
three charges per day, was approximately as follows:
1 foreman at $3 $3.00
5 furnace crews at $9.90 49.50
Unloading 21 tons of coal at 6c. 1.26
Loading 14 tons lead at 15c. 2.10
“ 7 tons gray slag at 15c. 1.05
——————
Total labor $56.91
21 tons coal at $2 $42.00
Flux and supplies 13.00
Blacksmithing and repairs 10.00
——————
Total $121.91
On the basis of 6.25 tons of wet ore, this would be $4.65 per ton. The
actual cost in seven consecutive months of 1900 was as follows: Labor,
$1.98 per ton; coal, $1.86; flux and supplies, $0.51; blacksmithing and
repairs, $0.39; miscellaneous, $0,017; total, $4.757. If the cost of
smelting the gray slag be reckoned at $8 per ton, and the proportion
of gray slag be reckoned at 0.25 ton per ton of galena concentrate,
the total cost of treatment of the latter comes to about $6.75 per ton
of wet charge, or about $7 per ton of dry charge. This cost could be
materially reduced in a larger and more perfectly designed plant.
The practice at Desloge did not compare unfavorably, either in respect
to metal extracted or in smelting cost, with the roast-reduction method
of smelting or the Scotch hearth method, as carried out in the plants
of similar capacity and approximately the same date of construction,
smelting the same class of ore, but the larger and more recent plants
in the vicinity of St. Louis could offer sufficiently better terms to
make it advisable to close down the Desloge plant and ship the ore to
them. One of the drawbacks of the reverberatory method of smelting
was the necessity of shipping away the gray slag, the quantity of
that product made in a small plant being insufficient to warrant the
operation of an independent shaft furnace.
PART III
SINTERING AND BRIQUETTING
THE DESULPHURIZATION OF SLIMES BY HEAP ROASTING AT BROKEN HILL[9]
BY E. J. HORWOOD
(August 22, 1903)
It is well known that, owing to the intimate mixture of the
constituents of the Broken Hill sulphide ores, a great deal of crushing
and grinding is required to detach the particles of galena from the
zinc blende and the gangue; and it will be understood, therefore, that
a considerable amount of the material is converted into a slime which
consists of minute but well-defined particles of all the constituents
of the ore, the relative proportions of which depend on the dual
characteristics of hardness and abundance of the various constituents.
An analysis of the slime shows the contents to be as follows;
Galena (PbS) 24.00
Blende (ZnS) 29.00
Pyrite (FeS₂) 3.38
Ferric oxide (Fe₂O₃) 4.17
Ferrous oxide (FeO) contained in garnets 1.03
Oxide of manganese (MnO) contained in rhodonite and garnets 6.66
Alumina (Al₂O₃) contained in kaolin and garnets 5.40
Lime (CaO) contained in garnets, etc. 3.40
Silica (SiO₂) 22.98
Silver (Ag) .06
——————
100.48
Galena, being the softest of these, is found in the slimes to a larger
extent than in the crude ore; it is also, for the same reason, in the
finest state of subdivision, as is well illustrated by the fact that
the last slime to settle in water is invariably much the richest in
lead, while the percentages of the harder constituents, zinc blende and
gangue, show a corresponding reduction in quantity, by reason of their
being generally in larger sized particles and consequently settling
earlier.
The fairly complete liberation of each of the constituent minerals
of the ore that takes place in sliming tends, of course, to help
the production of a high-grade concentrate by the use of tables and
vanners, and undoubtedly a fair recovery of lead is quite possible,
even with existing machines, in the treatment of fine slimes; but,
owing to the great reduction in the capacity of the machines, which
takes place when it is attempted to carry the vanning of the finer
slimes too far, and the consequently greatly increased area of the
machines that would be necessary, the operation, sooner or later,
becomes unprofitable.
The extent to which the vanner treatment of slimes should be carried
is, of course, less in the case of those mines owning smelters than
with those which have to depend on the sale of concentrates as their
sole source of profit. In the case of the Proprietary Company,
all slime produced in crushing is passed over the machines after
classification. A high recovery of lead in the form of concentrates
is, of course, neither expected nor obtained, for reasons already
explained; but the finest lead-bearing slimes are allowed to unite
with the tailings, which are collected from groups of machines, and
are then run into pointed boxes, where, with the aid of hydraulic
classification, the fine rich slimes are washed out and carried to
settling bins and tanks, where the water is stilled and allowed to
deposit its slime, and pass over a wide overflow as clear water. The
slime thus recovered amounts to over 1200 tons weekly, or about 11 per
cent., by weight, of the ore, and assays about 20 per cent. lead, 17
per cent. zinc, and 18 oz. silver, and represents, in lead value, about
11 per cent. of the original lead contents of the crude ore and rather
more than that percentage in silver contents. These slimes are thus a
by-product of the mills, and their production is unavoidable; but as
they are not chargeable with the cost of milling, they are an asset of
considerable value, more especially so since it has been demonstrated
that they can be desulphurized sufficiently for smelting purposes by a
simple operation, and, at the same time, converted into such a physical
condition as renders the material well suited for smelting, owing to
its ability to resist pressure in the furnaces.
The Broken Hill Proprietary Company has many thousands of tons of
these slimes which the smelters have hitherto been unable to cope with,
owing to the roasters being fully occupied with the more valuable
concentrates. Moreover, the desulphurization of slimes in Ropp
mechanical roasters is objectionable for various reasons, namely, owing
to the large amount of dust created with such fine material, resulting
injuriously to the men employed; also on account of the reduction in
the capacity of the roasters, and consequent increase in working cost,
owing to the lightness of the slime, especially when hot, as compared
with concentrates, and the necessity for limiting the thickness of
material on the bed of the roasters to a certain small maximum.
Further, the desulphurization of the slimes is no more complete with
the mechanical roasters than in the case of heap roasting, and the
combined cost of roasting and briquetting being quite three shillings
(or 75c.) per ton in excess of the cost of heap roasting, the
latter possesses many advantages. These heaps are being dealt with,
preparatory to roasting, by picking down the material in lumps of about
5 in. in thickness, while the fine dry smalls, unavoidably produced,
are worked up in a pug mill with water, and dealt with in the same way
as the wet slime produced from current work.
The slime, as produced by the mills, is run from bins into railway
trucks in a semi-fluid condition, and shortly after being tipped
alongside one of the various sidings on the mine is in a fit condition
to be cut with shovels into rough bricks, which dry with fair rapidity,
and when required for roasting are easily reloaded into railway trucks.
As each man can cut about 20 tons of bricks per day, the cost is small.
Various other methods of lumping the slime were tried, including
trucking the semi-fluid material on movable trams, alongside which were
set laths, about 9 in. apart, which enabled long slabs to be formed
9 in. wide and 5 in. thick, which were, after drying, picked up in
suitable lumps and loaded in platform trucks, thence on railway trucks.
Owing to the inferior roasting that takes place with bricks having flat
sides, which are liable to come into close contact in roasting, and
to the rather high labor cost, this method was discontinued. Another
method was to allow the slime to dry partially after being emptied
from railway trucks, and to break it into lumps by means of picks;
but this method entailed the making of an increased amount of smalls,
besides taking up more siding room, owing to the extra time required
for drying, as compared with the method now in use. Ordinary bricking
machines could, of course, be used, but when the cost of handling the
slime before and after bricking is counted, the cost would be greater
than with the simple method now in use; the material being in too
fluid a condition for making into bricks until some time elapses for
drying, a double handling would be necessitated before sending it to
the bricking machine. If, however, the slime could be allowed time to
dry sufficiently in the trucks, bricking by machinery would probably be
preferable. Rather more than 10 per cent. of smalls is made in handling
the lumps in and out of the railway trucks, and this is, as already
noted, worked up with water in a pug mill at the sintering works, and
used partly for covering the heaps with slime to exclude an excessive
amount of air. The balance is thrown out and cut into bricks, as
already described.
At the heaps the lumps are at present being thrown from one man to
another to reach their destination in the heap, but the sidings have
been laid out in duplicate with a view to enabling traveling cranes to
be used on the line next the heap, the lumps to be loaded primarily
into wooden skips fitting the trucks. It is probable, however, that
the lumps will require to be handled out of the skips into their place
in the heap, as the brittle nature of the material may be found to
render automatic tipping impracticable. A considerable saving in labor
would nevertheless accompany the use of cranes, which would likewise be
advantageous in loading the sintered material.
In order to reduce the inconvenience arising from fumes, length is very
desirable in siding accommodation, so that heap building may be carried
on at a sufficient distance from the burning kilns. It is for the same
reason preferable to build in a large tonnage at one time, lighting
the heaps altogether. As the heaps burn about two weeks only, long
intervals intervene, during which the fumes are absent.
In the experimental stages of slime roasting, fuel, chiefly wood, was
used in quantities up to 5 per cent., and was placed on the ground at
the bottom of the heap, where also a number of flues, loosely built
bricks, were placed for the circulation of air. The amount of fuel
used has, however, been gradually reduced, until the present practice
of placing no fuel whatever in the bottom was arrived at; but instead
less than 1 per cent. of wood is now burned in small enlargements of
the flues, under the outer portion of the pile, and placed about 12
ft. apart at the centers. This is found to be sufficient to start the
roasting operation within 24 hours of lighting, after which no further
fuel is necessary.
As regards the dimensions of the heaps, the width found most suitable
is 22 ft. at the base, the sides sloping up rather flatter than one to
one, with a flat section on top reaching about 7 ft. in hight. As there
is always about 6 in. of the outer crust imperfectly roasted, it is
advisable to make the length as great as possible, thus minimizing the
surface exposed. The company is building heaps up to 2000 ft. long.
During roasting care is required to regulate the air supply, the object
being to avoid too fierce a roast, which tends to sinter and partially
fuse the material on the outer portions of the lumps, while inside
there is raw slime. By extending the roast over a longer period this is
avoided, and a more complete desulphurization is effected. Experiments
conducted by Mr. Bradford, the chief assayer, demonstrated that, at a
temperature of 400 deg. C., the sulphide slime is converted into basic
sulphate, while at a temperature of 800 deg. C. the material becomes
sintered owing to the decomposition of the basic sulphate and the
formation of fusible silicate of lead.
In practice, the sulphur contents of the material, which originally
are about 14 per cent., become reduced to from 6.5 to 8.5 per cent.,
half in the form of basic sulphate and half as sulphides; much of the
material sinters and becomes matted together in a fairly solid mass.
The heaps are built without chimneys of any kind; a strip about 5
ft. wide along the crest of the pile is left uncovered by plastered
slime, and this, together with the open way in which the lumps are
built in, allows a natural draft to be set up, which can be regulated
by partly closing the open ends of the flues at the base of the pile.
Masonry kilns were used in the earlier stages with good results, which,
however, were not so much better than those obtained by the heap method
as to justify the expense of building, taking into consideration, too,
the extra cost of handling the roasted material in the necessarily more
confined space.
Much interest has been taken in the chemical reactions which take
place in the operation of desulphurization of these slimes, it being
contended, on the one hand, that the unexpectedly rapid roast which
takes place may be due to the sulphide being in a very fine state of
subdivision, and more or less porous, thus allowing the air ready
access to the sulphur, producing sulphurous acid gas (SO₂). On the
other hand, others, of whom Mr. Carmichael is the chief exponent, claim
that several reactions take place during the operation, connected
with the rhodonite and lime compounds present in the slimes, which he
describes as follows:
“The temperature of the kilns having reached a dull red heat, the
rhodonite (silicate of manganese) is converted into manganous oxide
and silica; at a rather higher temperature the calcium compounds are
also split up, with formation of calcium sulphide, the sulphur being
provided by the slimes. The air permeating the mass oxidizes the
manganese oxide and calcium sulphide into manganese tetroxide and
calcium sulphate respectively, as shown as follows;
3MnO + O = Mn₃O₄
CaS + 4O = CaSO₄,
and, as such, are carriers of a form of concentrated oxygen to the
sulphide slimes, with a corresponding reduction to manganous oxide and
calcium sulphide, as shown by the following equation, in the case of
lead:
PbS + 4Mn₃O₄ = PbSO₄ + 12MnO
PbS + CaSO₄ = PbSO₄ + CaS.
The oxidation of the manganous oxide and calcium sulphide is repeated,
and these alternate reactions recur until the desulphurization ceases,
or the kiln cools down to a temperature below which oxidation cannot
occur. These reactions, being heat-producing, provide part of the heat
necessary for desulphurization, which is brought about by certain
concurrent reactions between metallic sulphates and sulphide.
“The first that probably occurs is that in which two equivalents of the
metallic sulphide react on one of the metallic sulphate with reduction
to the metal, metallic sulphide, and sulphurous acid, as shown by the
following equation in the case of lead:
2PbS + PbSO₄ = 2Pb + PbS + 2SO₂.
“The metal so formed, in the presence of air, is oxidized, and in this
state reacts on a further portion of the metallic sulphide produced,
with an increased formation of metal and evolution of sulphurous acid,
according to the following equation, in the case of lead:
2PbO + PbS = Pb + SO₂.
“The metal so produced in this reaction is wholly reoxidized by the
oxygen of the air current, and being free to react on still further
portions of the metallic sulphide, repeats the reaction, and becomes
an important factor in the desulphurizing of the undecomposed portion
of the material. As the desulphurization proceeds, and the sulphate of
metal accumulates, reactions are set up between the metallic sulphide
and different multiple proportions of the metallic sulphate, with the
formation of metal, metallic oxide, and evolution of sulphurous acid,
as follows:
“With two equivalents of metallic sulphate to one equivalent of
metallic sulphide, in the case of lead, according to the following
equation:
PbS + 2PbSO₄ = 2PbO + Pb + 3SO₂.
“With three equivalents of metallic sulphate to one of metallic
sulphide, in the case of lead, according to the following equation:
PbS + 3PbSO₄ = 4PbO + 4SO₂.”
The volatility of sulphide of lead—especially in the presence of an
inert gas such as sulphurous acid—being greater than that of the
sulphate, oxide, or the metal itself, it might be thought that the
conditions are conducive to a serious loss of lead. This, however, is
reduced to a minimum, owing to the easily volatilized sulphide being
trapped, as non-volatile sulphate, by small portions of sulphuric
anhydride (SO₃), which is formed by a catalytic reaction set up
between the hot ore, sulphurous acid, and the air passing through
the mass. Owing to the non-volatility of the silver compounds in the
slimes, the loss of this metal has been found to be inappreciable. The
zinc contents of the slime are reduced appreciably, thus rendering the
material more suitable for smelting. After desulphurization ceases,
a few days are allowed for cooling off. On the breaking up of the
mass for despatch to the smelters, as much of the lower portion of
the walls is left intact as possible, so that it can be utilized for
the next roast, thus avoiding the re-building of the whole of the
walls.[10]
THE PREPARATION OF FINE MATERIAL FOR SMELTING
BY T. J. GREENWAY
(January 12, 1905)
In the course of smelting, at the works of the company known as the
Broken Hill Proprietary Block 14, material which consisted chiefly of
silver-lead concentrate and slime, resulting from the concentration
of the Broken Hill complex sulphide ore, I had to contend with all
the troubles which attend the treatment of large quantities of finely
divided material in blast furnaces. With the view of avoiding these
troubles, I experimented with various briquetting processes; and,
after a number of more or less unsatisfactory experiences, I adopted a
procedure similar to that followed in manufacturing ordinary bricks by
what is known as the semi-dry brick-pressing process. This method of
briquetting not only converts the finely divided material cheaply and
effectively into hard semi-fused lumps, which are especially suitable
for the heavy furnace burdens required by modern smelting practice, but
also eliminates sulphur, arsenic, etc., to a great extent; therefore,
it is capable of wide application in dealing with concentrate, slime,
and other finely divided material containing lead, copper and the
precious metals.
This briquetting process comprises the following series of operations:
1. Mixing the finely divided material with water and newly slaked lime.
2. Pressing the mixture into blocks of the size and shape of ordinary
bricks.
3. Stacking the briquettes in suitably covered kilns.
4. Burning the briquettes, so as to harden them, without melting, at
the same time eliminating sulphur, arsenic, etc.
1. The material is dumped into a mixing plant, together with such
proportions of screened slaked lime (usually from three to five per
cent.) and water as shall produce a powdery mixture which will, on
being squeezed in the hand, cohere into dry lumps. In preparing the
mixture, it is well to mix sandy material with suitable proportions
of fine, such as slime, in order that the finer material may act as a
binding agent.
The mixer used by me consists of an iron trough, about 8 ft. long,
traversed by a pair of revolving shafts, carrying a series of knives
arranged screw-fashion; and so placed that the knives on one shaft
travel through the spaces between the knives on the other shaft.
The various materials are dumped into one end of the mixing trough,
from barrows or trucks, and are delivered continuously at the other
end of the trough, into an elevator which conveys the mixture to the
brick-pressing plant.
2. The plant employed was the semi-dry brick-press. This machine
receives the mixture from the elevators, and delivers it in the form
of briquettes, which can at once be stacked in the kilns. It was found
that such material as concentrate and slime has comparatively little
mobility in the dies during the pressing operation; this necessitates
the use of a device which provides for the accurate filling of the
dies. It was also found that the materials treated by smelters vary
in compressibility, and this renders necessary the adoption of a
brick-pressing plant having plungers which are forced into the dies by
means of adjustable springs, brick-presses having plungers actuated by
rigid mechanism being extremely liable to jam and break.
3. Briquettes made from such material as concentrate and slime vary
in fusibility; they are also combustible, and while being burned they
produce large quantities of smoke containing sulphurous acid and other
objectionable fumes. It is therefore necessary that such briquettes be
burned in kilns provided with arrangements for accurately controlling
the burning operations, and for conveniently disposing of the smoke.
Suitable kilns, which will contain from 30 to 50 tons of briquettes
per setting, are employed for this purpose. Regenerative kilns of the
Hoffman type might be used for dealing with some classes of material,
but, for general purposes, the kilns as designed here will be found
more convenient.
The briquettes are stacked according to the character of the material
and the object to be obtained. The various methods of stacking, and the
reasons for adopting them, can be readily learned by studying ordinary
brick-burning operations in any large brick-yard. After the stacking
is complete the kiln-fronts are built up with burnt briquettes produced
in conducting previous operations, and all the joints are well luted.
4. In burning briquettes made from pyrite or other self-burning
material, it is simply necessary to maintain a fire in the kiln
fireplaces for a period of from 10 to 20 hours. When it is judged that
this firing has been continued long enough, the fire-bars are drawn
and the fronts are luted with burnt briquettes in the same manner as
the kiln-fronts. Holes about two inches square are then made in these
lutings, through which the air required for the further burning of the
briquettes is allowed to enter the kilns under proper control. After
the fireplaces are thus closed the progress of the burning, which
continues for periods of from three to six days, is watched through
small inspection holes made in the kiln-fronts; and when it is seen
that the burning is complete the fronts are partially torn away,
in order to accelerate the cooling of the burnt briquettes, which
are broken down and conveyed to the smelters as soon as they can be
conveniently handled.
When briquettes made from pyrite concentrate, or of other free-burning
material, are thus treated, they are not only sintered but they are
also more or less effectively roasted, and it may be taken for granted
that any ore which can be effectively roasted in the lump form in kilns
or stalls will form briquettes that will both sinter and roast well;
indeed, one may say more than this, for briquettes which will sinter
and roast well can be made from many classes of ore that cannot be
effectively treated by ordinary kiln-and stall-roasting operations;
and, moreover, good-burning briquettes may be made from mixtures of
free-burning and poor-burning material. Briquettes containing large
proportions of pyrite or other free-burning material will, unless the
air-supply is properly controlled, often heat up to such an extent as
to fuse into solid masses, much in the same manner as matte of pyritic
ore will melt when it is unskilfully handled in roasting. In dealing
with material which will not burn freely, such as roasted concentrate,
the briquetting is conducted with the intention of sintering the
material; and in this case the firing of the kilns is continued for
periods of from three to four days, the procedure being similar in
every way to that followed in burning ordinary bricks.
When conducting my earlier briquetting operations I made the
briquettes by simply pugging the finely divided material, following
a practice similar to that adopted in producing “slop-made” bricks
by hand. This method of making the briquettes was attended with a
number of obvious disadvantages, and was abandoned as soon as the
semi-dry brick-pressing plant became available. The extent to which
this process, or modifications of it, may be applied is shown by the
fact that, following upon information given by me, the Broken Hill
Proprietary Company adopted a similar method of sintering and roasting
slime, consisting of about 20 per cent. galena, 20 per cent. blende,
and 60 per cent. silicious gangue. The procedure followed in this
case consisted of simply pugging the slime, and running the pug upon
a floor to dry; afterward cutting the dried material into lumps by
means of suitable cutting tools, and then piling the lumps over firing
foundations, following a practice similar to that pursued in conducting
ordinary heap-roasting. This company is now treating from 500 to 1000
tons of slime weekly in this manner. It is, however, certain that
better results would attend the treatment of this material by making
this slime into briquettes and burning them in kilns.
The cost of briquetting and burning material in the manner first
described, with labor at 25c. per hour, and wood or coal at $4 per ton,
amounts to from $1 to $1.50 per ton of material.
THE BRIQUETTING OF MINERALS
BY ROBERT SCHORR
(November 22, 1902)
The value of briquetting in connection with metallurgical processes and
the manufacture of artificial stone is well understood and appreciated.
In smelting plants there is always more or less flue dust, fine ores,
and sometimes fine concentrates to be treated, but the charging
of such fine material directly into a furnace would cause trouble
and irregularities, and would lessen its capacity also. As mineral
briquetting cannot be effected without considerable wear upon the
machinery and without quite appreciable expense in binder, labor, and
handling, many smelters try to avoid it.
The financial question, however, is not as serious as it may at first
appear, and taking the large output of modern briquetting machines in
consideration, the cost for repairs amounts only to a few cents per ton
of briquetted material. The total cost depends in the first place on
the cost of labor, power and the binder, and in most American smelters
it varies between $0.65 and $1.25 per ton of briquettes.
Ordinary brick presses, with clay as a binder, were used in Europe as
well as in this country, but they are too slow and expensive for large
propositions and the presence of clay is usually undesirable.
The English Yeadon (fuel) press has also been used for some years at
the Carlton Iron Company’s Works at Ferryhill in England, and at the
Ore and Fuel Company’s plant at Coatbridge in the same country; also by
some Continental firms. Dupuis & Sons, Paris, furnished a few presses
which are mostly used for manganese and iron ores and pyrites. In
some localities coke dust is added. The making of clay briquettes or
mud-cakes is the crudest form of briquetting; but while heat has to
be expended to evaporate the 40 to 50 per cent. of moisture in them,
and while considerable flue dust is made, this method is better than
feeding fine ore or flue dust directly into the furnace.
The only other method of avoiding briquetting is by fusing ore fines in
slagging reverberatory furnaces and by adding flue dust in the slagging
pit, thus incorporating it with the slagging ore. This is practised
sometimes in silver-lead smelters, but in connection with copper or
iron smelters it is not practicable.
In briquetting minerals a thorough mixing and kneading is of the first
importance. If this is done properly a comparatively low pressure will
suffice to create a good and solid briquette, which after six to eight
hours of air-drying, or after a speedier elimination of the surplus of
moisture in hot-air chambers, will be ready for the furnace charge. A
good briquette should permit transportation without excessive breakage
or dust a few hours after being made, and it should retain its shape in
the furnace until completely fused, so as to create as little flue dust
as possible. The briquette should be dense, otherwise it will crumble
under the influence of bad weather.
The two presses on the American machinery market are the type built by
the Chisholm, Boyd & White Company, of Chicago, and the briquetting
machine manufactured by the H. S. Mould Company, of Pittsburg. Both are
extensively used, and in many metallurgical plants it will pay well to
adopt them.
From 4 to 6 per cent. of milk of lime is generally used as binder,
and this has a desirable fluxing influence also. A complete outfit
comprises, besides the press, a mixer for slacking the lime, and a
feed-pump which discharges the liquid in proportion into the main mixer
wherein the ore fines, flue dust, or concentrates are shoveled.
The Chisholm, Boyd & White Company’s press makes 80 briquettes per
minute, which, with a new disk, are of 4 in. diameter and 2½ in. hight,
thus giving about 872 cu. ft. of briquette volume per 10 hours, or 50
to 80 tons, depending on the weight of the material. With the wear of
the disk the hight of the briquettes is reduced and consequently the
capacity of the machine also. The disk weighs about 1600 lb., and as
most large smelters have their own foundries it can be replaced with
little expense. About 30 effective horse-power is usually provided for
driving the apparatus. The machine is too well known to metallurgists
and engineers to require further comment or description.
The H. S. Mould Company has also succeeded in making its machine a
thorough practical success. This machine is a plunger-type press. The
largest press built employs six plungers, and at 25 revolutions it
makes 150 briquettes of 3 in. diameter and 3 in. hight, or 1080 cu. ft.
per 10 hours. Its rated capacity is 100 tons per 10 hours.
In using a plunger-type press the material should not contain more
than 7 per cent. mechanical moisture. If wet concentrates have to
be briquetted it is necessary to add dry ore fines or flue dust to
arrive at a proper consistency. The briquettes are very solid and only
air-drying for a few hours is necessary.
The cylindrical shape of briquettes is very good, as it insures
a proper air circulation in the furnace and consequently a rapid
oxidation and fusion.
The wear of the Mould Company’s press is mostly confined to the chilled
iron bushings and to the pistons. Auxiliary machinery consists of
the slacker, the feeder and the main mixer. The press is of a very
substantial design, and it is claimed that the cost of repairs does not
amount to more than 3c. per ton of briquettes.
Wear and tear is unavoidable in a crude operation like briquetting; to
treat flue dust, ore fines, and fine concentrates successfully, it is
almost absolutely necessary to resort to it.
Edison used a number of intermittent-acting presses at his magnetic
iron-separation works in New Jersey, but this plant shut down some time
ago.
A BRICKING PLANT FOR FLUE DUST AND FINE ORES
BY JAMES C. BENNETT
(September 15, 1904)
The plant, which is here described, for bricking fine ores and flue
dust, was designed and the plans produced in the engineering department
of the Selby smelter. The machinery contained in the plant consists of
a Boyd four-mold brick press, a 7 ft. wet pan or Chile mill, a 50 h.p.
induction motor, and a conveyor-elevator, together with the necessary
pulleys and shafting.
The press, Chile mill, and motor need no special mention, as they all
are from standard patterns and bought, without alterations, from the
respective builders. The Chile mill was purchased from the builders
of the brick press. The conveyor-elevator was built on the premises
and consists of a 14 in. eight-ply rubber belt, with buckets of sheet
steel placed at intervals of 6 in., running over flanged pulleys. The
buckets, or more properly speaking the flights, are made from No.
12 steel plate, flanged to produce the back and ends, with the ends
secured to the flanged bottom by one rivet in each. The plant has been
in operation for sixteen months and there have been few or no repairs
to the elevator, except to renew the belt, which is attacked by the
acid contained in the charges. This first belt was in continuous use
for nine months. As originally designed, the capacity was 100 tons per
day of 12 hours, but this was found to require a speed so high that
the workmen were unable to handle the output of the press. The speed
was, consequently, reduced about 25 per cent., which brings the output
down to about 75 tons per day. This output, as expressed in weight,
naturally varies somewhat owing to the variation in the weight of the
material handled.
It is probable that the capacity could be increased to about 90 tons
by enlarging the bricks, which could be done, but would require a
considerable amount of alteration in the machine, as it is designed to
produce a standard sized building brick. By this method of increase,
however, the work of handling would not be materially increased,
because the number of bricks would be the same as with the present
output of 75 tons; there would be about 16 per cent. more to handle,
by weight. Working on the basis of 100 tons capacity, the bins were
designed to afford storage room for about three days’ run, or a little
over 300 tons. The bins are made entirely of steel, in order that
the hot material may be dumped into them directly from the roasting
furnaces, thus saving one handling. In order that there may be room
for several kinds of material, the bins are divided into seven
compartments, three on one side and four on the other. The lower part
is of ⅜ in. steel plate, and the upper, about one-half the hight, of
5/16 in. plate.
It may be well to call attention to the method of handling the
material, preparatory to its delivery to the brick press. The bins are
constructed, as will be seen by the drawing, with their floor set 2.5
ft. above the working floor, which enables the workmen to reach the
material with a minimum effort. The floor of the bins project 2.5 ft.
in front of the face, thus forming a platform on which the shoveling
may be done without the necessity of bending over. In this projecting
platform are cut rectangular holes 12 × 18 in., which are placed
midway between the openings in the front of the bins and furnished
with screens to stop any stray bolts or other coarse material that
might injure the press. This position of the holes through the platform
was adopted so that, in the event of the material running out beyond
the opening in the face, it would not fall directly upon the floor.
Two buckets are provided, with a capacity of 7 cu. ft. each, which is
the size of a single charge of the Chile mill. These buckets have a
hopper-shaped bottom fixed with a swinging gate which is operated by
the foot; thus the bucket can be run over the pan of the Chile mill and
the charge dumped directly into it. The buckets run on an overhead iron
track (1 in. by 3 in.) hung 7 ft. in the clear, above the floor.
The method of making up the charge is as follows: The bucket is
run under the hole in the platform nearest to the compartment
containing the material of which the charge is partly composed, and
a predetermined number of shovelfuls is drawn out and put into the
bucket, which is then pushed on to the next compartment from which
material is wanted, where the operation is repeated. After charging
into the bucket the requisite amount of ore or flue dust, the bucket
is run to the back of the building, where the necessary amount of lime
(slaked) is added. By putting the lime in last, it is so surrounded by
the dust or ore that it has not the opportunity to stick to the sides
of the bucket in discharging, as it otherwise would.
[Illustration: FIG. 1 (_a_).—Plant for Bricking Ores, Selby Smelter.
(Plan.)]
The number of men required to operate the entire plant, exclusive
of those employed in bringing the material to the bins and emptying
the cars into them, is 12, placed as follows; One preparing the lime
for use, one removing the charge from the mill and supplying the
elevator-conveyor, which is accomplished by means of a specially
shaped, long-handled shovel; one keeping the supply spout of the press
clear (an attempt was made to do this mechanically, but was found to be
unsuccessful, owing to the extremely sticky nature of the material, and
so was discarded in favor of manual labor); one to control the press in
case of mishap and to keep the dies clean; one oiler; three receiving
the bricks from the press and taking the brick-loaded cars from the
press to the drying-house, and two placing the bricks on the shelves.
[Illustration: FIG. 1 (_b_).—Plant for Bricking Ores, Selby Smelter.
(Elevation.)]
The drying-house scarcely requires description; it is but a roofed
shed, without sides, fitted with stalls into which the bricks are set
on portable shelves, as close as working conditions will permit. The
means of drying, at the present time, is by the natural circulation
of air, but a mechanical system is in contemplation, by which the
air will be drawn into the building from the outside and forced to
find its way out through the bricks. The drying-house is adjacent to
the pressing plant, in fact forms the back of it, so that there is a
minimum distance to haul the product. The time required for drying the
bricks sufficiently for them to withstand the necessary handling is,
depending on the weather, from two to eight days, the usual time being
about three days.
PART IV
SMELTING IN THE BLAST FURNACE
MODERN SILVER-LEAD SMELTING[11]
BY ARTHUR S. DWIGHT
(January 10, 1903)
The rectangular silver-lead blast furnace developed in the Rocky
Mountains has an area of 42 × 120 to 48 × 160 in. at the tuyeres; 54
× 132 to 84 × 200 in. at the top; and hight from tuyere level to top
of charge of 15 to 21 ft. Such a furnace smelts 80 to 200 tons of
charge (ore and flux, but not slag and coke) per 24 hours. The slag
that has to be resmelted amounts to 20 to 60 per cent. of the charge.
Coke consumption is 12 to 16 per cent. of the charge. The blast
pressure ranges from 1.5 to 4 lb. per square inch, averaging close to
2 lb. Gases of hand-charged furnaces are taken off through an opening
below the charge-floor, the furnace being fed through a slot (about
20 in. wide, extending nearly the whole length of the furnace) in the
iron floor-plates; or through a hood (brick or sheet iron) above the
charge-floor level, with a down-take to the flues, charge-doors being
provided on each side of the hood, extending preferably the whole
length of the furnace and usually having a sill a few inches high which
compels the feeder to lift his shovel.
When a silver-lead blast furnace is operating satisfactorily, the
following conditions should obtain; (1) A large proportion of the lead
in the charge should appear as direct bullion-product at the lead-well.
(2) The slag should be fluid and clean. (3) The matte should be low
in lead. (4) The furnace should be cool and quiet on top, making a
minimum quantity of lead-fume and flue-dust, and the charges should
descend uniformly over the whole area of the shaft. (5) The furnace
speed should be good. (6) The furnace should be free from serious
accretions and crusts; that is to say, the tuyeres should be reasonably
bright and open, and the level of the lead in the lead-well should
respond promptly to variations of pressure, caused by the blast and by
the hight of the column of molten slag and matte inside the furnace—an
indication that ample connection exists between the smelting column and
the crucible. Good reduction (using that term to express the degree in
which the furnace is manifesting its reducing action) is obtained when
the first three of the above conditions are satisfied.
For any given furnace there are five prime factors, the resultant of
which determines the reduction, namely: (_a_) Chemical composition
of the furnace charges; (_b_) proportion and character of fuel;
(_c_) air-volume and pressure, to which might perhaps also be added
temperature of blast; for, although hot blast has not yet been
successfully applied in lead-smelting practice, I believe it is only
a question of time when it will be; (_d_) dimensions and proportions
of smelting furnace; (_e_) mechanical character and arrangement of the
smelting column.
All but one of the above factors can be intelligently gaged. The
mechanical factor, however, can be expressed only in generalities and
indefinite terms. A wise selection of ores and proper preliminary
preparation, crushing the coarse and briquetting the fine, will do
much to regulate it, but all this care may be largely nullified by
careless feeding. The importance and possibilities of the mechanical
factor are generally overlooked and its symptoms are wrongly diagnosed.
For instance, the importance of slag-types has undoubtedly been
considerably exaggerated at the expense of the mechanical factor.
Slags seldom come down exactly as figured. We must know our ores and
apply certain empirical corrections to the iron, sulphur, etc., based
on previous experience with the ores; but these empirical corrections
may represent also an unformulated expression of the influence of the
mechanical factor on the reduction—a function, therefore, of the ruling
physical complexion of the ores, and the peculiarities of the feeding
habitually maintained in the works concerned. With a given ore-charge
large reciprocal variations may be produced in the composition of
slag and matte by merely changing the mechanical conditions of the
smelting column, and since the efficient utilization of both fuel and
blast must be controlled in the same way, the mechanical factor may be
considered, perhaps, the dominating agent of reduction. Inasmuch as
there is no way of gaging it, however, the only recourse is to seek a
correct adjustment and maintain it as a positive constant, after which
slag, fuel and blast may be with much greater certainty adjusted toward
efficiency of furnace work and metal-saving.
_Behavior of Iron._—The output of lead is so dependent upon the
reactions of the iron in the charge that the chief attention may well
be fixed upon that metal as the key to the situation. The success of
the process depends largely upon reducing just the right amount of
iron to throw the lead out of the matte, the remainder of the iron
being reduced only to ferrous oxide and entering the slag. Too much
iron reduced will form a sow in the hearth. Iron is reduced from its
oxides principally by contact with solid incandescent carbon, and by
the action of hot carbon monoxide. Reduction by solid carbon is the
more wasteful, but there is in lead smelting an even more serious
objection to permitting the reduction to be accomplished by that means,
which leads to comparatively hot top and more or less volatilization of
lead. Reduction by carbon monoxide is the ideal condition for the lead
furnace. It means keeping the zone of incandescence low in the charge
column, leaving plenty of room above for the gases to yield up their
heat to, and exercise their reducing power on, the descending charge,
so that by the time they escape they will be well-nigh spent. Their
volume and temperature will be diminished, and the low velocity of
their exit will tend to minimize the loss of lead in fume and flue dust.
The idea that high temperatures in lead blast furnaces should be
avoided is based on a misconception. Temperatures must exist which
are sufficiently high to volatilize all the lead in the charge, if
other conditions permit. A high temperature before the tuyeres means
fast smelting; and fast smelting, under proper conditions, means a
shortening of the time during which the lead is subject to scorifying
and volatilizing influences. A rapidly descending charge, constantly
replenished with cold ore from above, absorbs effectively the heat of
the gases and acts as a most efficient dust and fume collector. In
considering long flues, bag-houses, etc., it should be kept in mind
that the most effective dust collector ought to be the furnace itself.
In the practice of twelve years ago and earlier, particularly when
using mixed coke and charcoal, reduction by carbon was probably the
rule; and the percentage of fuel required was very high. There is good
reason to think we have still much room for improvement along this line
in our average practice of today.
_Volume of Blast._—It is customary to supply a battery of furnaces
from a large blast main, connected with a number of blowers. Inasmuch
as the air will take preferably the line of least resistance, if the
internal resistance of any one furnace be increased the volume of air
it will take will be diminished and the others will be favored unduly.
Only by keeping all the furnaces on approximately the same charge, with
the same hight of smelting column, can anything like uniformity of
operation and close regulation be secured. The rational plan would seem
to be to have a separate blower, of variable speed, directly connected
to each furnace, but this plan, which has had a number of trials, has
usually been abandoned in favor of the common blast main. Trials by
myself, extending over considerable periods, have been so uniformly
favorable, however, that I am forced to ascribe the failure of others
to some outside reason.
The peculiar atmosphere required in the lead blast furnace depends
upon the correct proportion of two counteractive elements, carbon and
oxygen. If given too much air the furnace will show signs of deficient
reduction, commonly interpreted as calling for more fuel, which will
be sheer waste since its object is to burn up surplus air. There will
be an additional waste through the extra coal burned under the steam
boilers. The true remedy would be to cut down the quantity of air.
Burning up excessive coke is as hard work as smelting ore. Too much
fuel invariably slows up a furnace; it also drives the fire upward and
gives predominance to reduction by solid carbon. The maintenance of a
minimum fuel percentage, with a correctly adjusted volume of air, will
tend to promote the conditions under which iron will be reduced by the
gases, rather than by solid carbon.
_Pressure of Blast._—Pressure necessarily involves resistance; and
the blast-pressure, as registered by a simple mercury-gage on the
bustle-pipe, may be increased in two ways: (1) By increasing the volume
of air forced through the interstices in the charge. This is the
wrong way; but, unfortunately, it is only too common in our practice,
and therefore deserves to be mentioned, if only to be condemned. (2)
By leaving the volume of air unchanged, but increasing the friction
offered by the interstitial channels, either by making them smaller in
aggregate cross-section (which means a finer charge), or by making them
longer (which means a higher smelting column). A correctly graduated
internal resistance is, therefore, the only true basis for a high blast
furnace, which, when so produced, will bring about rapid smelting, a
low zone of incandescence, and a very vigorous action upon the ores by
the gases in their retarded ascent through the charge column. These
conditions promote the reduction of iron by CO. The adjustment of
internal resistance, which is thus clearly the main factor, can be
accomplished only by the correct feeding of the furnace.
_Feeding the Charge._—It is self-evident that, the more thorough the
preliminary preparation of the charge before it reaches the zone of
fusion, the more rapidly can the actual smelting proceed. A piece of
raw ore that finds itself prematurely at the tuyeres, without having
been subjected to the usual preparatory processes of drying, heating,
reduction, etc., must remain there until it is gradually dissolved or
carried away mechanically in the slag. Any such occurrence must greatly
retard the process. It would seem, by the same reasoning, that an
intimate mixture of the ingredients of the charge should expedite the
smelting, and I advocate the intimate mixture of the charge ingredients
in all cases.
The theory of feeding is simple, but not so the practice. If the
charge column were composed of pieces of uniform size, the ascending
gases would find the channel of least resistance close to the furnace
walls and would take it preferably to the center of the shaft. The
more restricted channel would necessitate a higher velocity, so that
not only would the center of the charge be deprived of the action of
the gases, but also the portion traversed would be overheated; many
particles of ore would be sintered to the walls or carried off as flue
dust; slag would form prematurely; fuel would be wasted; in short,
all the irregularities and losses which accompany over-fire would be
experienced. In practice the charge is never uniform, but is a mixture
of coarse and fine. By lodging the finer material close to the walls
and placing the coarser in the center, an adjustment may be made which
will cause the gases to ascend uniformly through the smelting column.
A furnace top smoking quietly and uniformly over its whole area is the
visible sign of a properly fed furnace.
_Effect of Large Charges._—It has frequently been remarked that,
within certain limits, large charges give more favorable results
than small ones; and numerous attempts have been made to account
for this fact. My observations lead me to offer the following as a
rational explanation—at least in cases where ore and fuel are charged
in alternate layers. Large ore-charges mean correspondingly large
fuel-charges. The gases can pass readily through the coke; and hence
each fuel-zone tends to equalize the gas currents by giving them
another opportunity to distribute themselves over the whole furnace
area, while each layer of ore subsequently encountered will blanket the
gases, and compel them to force a passage under pressure, which is the
manner most favorable to effective chemical action.
In mechanically fed furnaces the charges of ore and fuel are usually
dropped in simultaneously from a car and the separate layers thus
obliterated, and the distributing zones which are such a safeguard
against the consequences of bad feeding are lacking, hence more care
must be exercised to secure proper placing of the coarse and fine
material. This may throw some light on the failure of most of the early
attempts at mechanical feeding.
_Mechanical Character of Charge._—Very fine charges blanket the gases
excessively and cause them to break through at a few points, forming
blow-holes, which seriously disturb the operation, cause loss of raw
ore in the slag, and are accompanied by all the evils of over-fire. A
charge containing a few massive pieces, the rest being fine, is a still
more unfavorable combination. A very coarse charge permits too ready an
exit to the gases, and in the end tends likewise to over-fire and poor
reduction. The remedy is to briquette the fine ore (though preferably
not all of it), and crush the coarse to such degree as to approach an
ideal result, which may be roughly described as a mixture in which
about one-third is composed of pieces of 5 to 2 in. in diameter,
one-third pieces of 2 to 0.5 in., and the remaining third from 0.5 in.
down. The coke is better for being somewhat broken up before charging,
and a reasonable amount of coke fines, such as usually accompanies
a good quality of coke, is not in the least detrimental. The common
practice of handling the coke by forks and throwing away the fines
is to be condemned as an unwarranted waste of good fuel. The slag on
the charge should be broken to pieces at most 6 in. in diameter. The
common practice of throwing in whole butts of slag-shells is bad.
There is no economy in using the slag hot; cold charges, not hot,
are what we want. A reasonable amount of moisture in the charge is
beneficial, providing it be in such form as to be readily dried out. It
is often advantageous to wet the ore mixtures while bedding them, or
to sprinkle the charges before feeding. The driving off of this water
must consume fuel, but not so much as if the smelting zone crept up.
Large doses of water applied directly to the furnace are unpardonable
under any circumstances, however, though they are sometimes indulged
in as a drastic measure to subdue excessive over-fire when other and
surer means are not recognized. One of the chief merits of moderate
sprinkling before charging is that it gives in many cases a more
favorable mechanical character, approximating a lumpy condition in too
fine a charge, and assisting to pack a too coarse one.
_Different Behavior of Coarse and Fine Ore._—In taking up a shovelful
of ore, the fine will be observed to predominate in the bottom and
center, and the coarse on the top and sides. When thrown from the
shovel, the coarse will outstrip the fine and fall beyond it. In making
a conical pile the coarse ore will roll to the base, leaving the fine
near the apex. This difference in the action of the mobile coarse ore
and the sluggish fines is the key to the practical side of feeding,
both manual and mechanical. It is not sufficient to tell the feeder to
throw the coarse in the middle and the fine against the sides; if it be
easier to do it some other way such instructions will count for little.
The desired result can be best secured by making the right way easier
than the wrong way.
It is generally conceded that the open-top furnaces, fed by hand
through a slot in the floor-plates, do not give as satisfactory results
as the hooded furnaces with long feed-doors on both sides. In the
open-top furnace it is comparatively difficult to throw to the sides;
the narrower the slot the greater the difficulty. The major part of the
charge will drop near the center, making that place higher than the
sides. The fine ore will tend to stay where it falls, while the coarse
will tend to roll to the sides, thus leading to an arrangement of the
charge just the reverse of what it ought to be. In the hooded furnace
most of the material will naturally fall near the doors, causing the
sides to be higher than the center toward which the coarse will roll,
while the force of the throw as the ore is shoveled in will also have
a tendency to concentrate the coarse material in the center.
Once a proper balance of conditions has been found, absolute
regularity of routine is the secret of good results. An experienced
and intelligent feeder owes his merit to his conscientious regularity
of work. He may have to vary his program somewhat when he encounters
a furnace that is suffering from the results of bad feeding by a
predecessor; but his guiding principle is first to restore regularity,
and then maintain it. A poor feeder can bring about, in a single
shift, disorders that will require many days to correct, if indeed
they are corrected at all during the campaign. The personal element is
productive of more harm than good.
_Mechanical Feeding._—If it be admitted that the work of a feeder
is the better the more it approximates the regularity of that of a
machine, it ought to be desirable to eliminate the personal factor
entirely and design a machine for the purpose, which would be a
comparatively simple matter if it be known just what we want to
accomplish. No valid ground now exists for prejudice against mechanical
feeding in lead smelting. It is in successful operation in a number
of large works, and is being installed in others. Our furnaces have
outgrown the shovel; we have passed the limit of efficiency of the
old methods of handling material for them. We must come to mechanical
feeding in spite of ourselves. But whatever may be the motive leading
to its introduction, its chief justification will be discovered,
after it has been successfully installed and correctly adjusted, in
the consequent great improvement of general operating results, metal
saving, etc. It will remove one of the most uncertain factors with
which the metallurgist has to deal, thereby bringing into clearer view
for study and regulation the other factors (fuel and blast proportion,
slag composition, etc.) in a way that has hardly been possible under
the irregularities consequent upon hand feeding.
MECHANICAL FEEDING OF SILVER-LEAD BLAST FURNACES[12]
BY ARTHUR S. DWIGHT
(January 17, 1903)
_Historical._—A silver-lead furnace fed by means of cup and cone was in
operation in 1888 at the works of the St. Louis Smelting and Refining
Company at St. Louis, Mo., but it is probable that previous attempts
had been made, since Hahn refers (“Mineral Resources of the United
States,” 1883) in a general way to experiments with this device, which
were unsuccessful because the heat crept up in the furnace and gave
over-fire. At the time of my visit to the St. Louis works (in 1888)
the furnaces were showing signs of over-fire, but this may not have
been their characteristic condition. A. F. Schneider, who built the St.
Louis furnaces, afterward erected, at the Guggenheim works at Perth
Amboy, N. J. , round furnaces with cup and cone feeders, but although
good results are said to have been obtained, the running of refinery
products is no criterion of what they would do on general ore smelting.
_Cup and Cone Feeders._—The cup and cone is an entirely rational device
for feeding a round furnace, but is quite unsuitable for feeding a
rectangular one. Furnaces of the latter type were installed for copper
smelting at Aguas Calientes, Mex., with two sets of circular cup and
cone feeders, but disastrous results followed the application of this
device to lead furnaces. The reason is clear when it is considered that
a circular distribution cannot possibly conform to the requirements
of a rectangular furnace. A more rational device was designed for the
works at Perth Amboy, N. J.
[Illustration: FIG. 2.—Perth Amboy, N. J. , Lead Furnace. Vertical
section at right angles to Fig. 3.]
_Pfort Curtain._—About ten years ago some of the American smelters
adopted the Pfort curtain, which, as adapted to their requirements,
consisted of a thimble of sheet iron hung from the iron deck plates so
as to leave about 15 in. of space between it and the furnace walls,
this space being connected with the down-take of the furnace. The
thimble was kept full of ore up to the charge-floor. This device was
popular for a time, chiefly because it prevented the furnace from
smoking and diminished the labor of feeding, but it was found to give
bad results in the furnaces, it being impossible to observe how the
charge sunk (except by dropping it below the thimble), while the
curtain had to be removed in order to bar down accretions, and, most
important, it caused irregular furnace work and high metal losses,
because it effected a distribution of the coarse and fine material
which was the reverse of correct, the evil being emphasized by the
taking off of the gases close to the furnace walls.
[Illustration: FIG. 3.—Perth Amboy, N.J., Lead Furnace. Vertical
section at right angles to Fig. 2.]
_Terhune Gratings._—R. H. Terhune designed a device (United States
patent No. 585,297, June 29, 1897), which comprised two grizzlies,
one on each side of the furnace, sloping downward from the edge of
the charge-floor toward the center line of the furnace. The bars
tapered toward the center of the furnace, the open spaces tapering
correspondingly toward the sides, so that as the charge was dumped on
them a classification of coarse and fine would be effected. This device
is correct in conception.
_Pueblo System._—In the remodeling of the plant of the Pueblo Smelting
and Refining Company in 1895, under the direction of W. W. Allen,
mechanical feeding was introduced, and the system was the first one to
be applied successfully on a large scale. The furnaces of this plant
are 60 × 120 in. at the tuyeres, with six tuyeres, 4 in. in diameter
on each side, the nozzles (water cooled) projecting 6 in. inside the
jackets. The hight of the smelting column above the tuyeres is 20 ft.
The gases are taken off below the charge-floor, and the furnace tops
are closed by hinged and counter-weighted doors of heavy sheet iron,
opened by the attendant, just previous to dumping the charge-car. In
the side walls of the shaft are iron door-frames, ordinarily bricked
up, but giving access to the shaft for repairs or barring out without
interfering with the movement of the charge-car. Extending across the
shaft, about 18 in. above the normal stock line, are three A-shaped
cast-iron deflectors, dividing the area of the shaft into four equal
rectangles.
The general arrangement of the plant is shown in Fig. 4. From the
charge-car pit there extends an inclined trestle, on an angle of 17
deg. to the charge-floor level, in line with the battery of furnaces.
The gage of the track is approximately equal to the length of the
furnaces at the top. The charge-car, actuated by a steel tail-rope,
moves sideways on this track from the charging-pit to any furnace
in the battery. The hoisting drums are located at the crest of the
incline, inside of the furnace building. At the far end of the latter
there is a tightener sheave, with a weight to keep proper tension on
the tail-rope. The charge-car has a capacity of 5 tons. It has an
A-shape bottom, and is so arranged that one attendant can quickly trip
the bolt and discharge the car.
[Illustration: FIG. 4.—Pueblo System. Longitudinal vertical section
through incline.]
While the car is making its trip the charge-wheelers are filling their
buggies, working in pairs, each man weighing up a halfcharge of a
particular ingredient. They then separate, each taking his proper place
in the line of wheelers on either side. When the car has returned, the
wheelers successively discharge their buggies into opposite ends of
the car. The coke is added last, to avoid crushing. The system is not
strictly economical of labor, since the wheelers, who must always be
ready for their car, have to wait for its return, which necessitates
more wheelers than would otherwise be required. Figs. 5, 6 and 7 show
the car.
[Illustration: FIG. 5.—Pueblo Charge-car. (Side elevation.)]
A vertical section through the car filled by dumping from the two ends
will show an arrangement of coarse and fine, which is far from regular.
Analyzing its structure, we shall find a conical pile near each end,
with a valley between them, in which coarse ore will predominate. The
deflectors in the furnace, previously referred to, serve to scatter
the fines as the charge is dropped in. Without them the feeding of the
furnace would be a failure; with them it is successful, though not so
completely as might be, the furnaces having a tendency to run with hot
tops. With the battery of seven furnaces, each smelting an average of
100 tons of ore per day, the saving, as compared with hand-feeding,
was $63 per day, or 9c. per ton of ore, this including cost of steam,
but not wear and tear on the machinery. This is distinctly a maximum
figure; with fewer furnaces the fixed charges of the mechanical feed
would soon increase the cost per ton to such a figure that the two
systems would be about equal in economy.
[Illustration: FIG. 6.—Pueblo Charge-car. (Plan.)]
[Illustration: FIG. 7.—Pueblo Charge-car. (End elevation.)]
_East Helena System._—This was introduced at the East Helena plant of
the United Smelting and Refining Company by H. W. Hixon. The plant
comprised four lead furnaces, each 48 × 136 in., with a 21 ft. smelting
column. They were all open-top furnaces, fed through a slot over the
center, the gases being taken off below the floor. They were capable of
smelting about 180 tons of charge (ore and flux) per 24 hours, using
a blast of 30 to 48 oz., furnished by two Allis duplex, horizontal,
piston blowers, air-cylinders 36 in. diam., 42 in. stroke, belted
from electric motors. The Hixon feed was designed to meet existing
conditions, without irrevocably cutting off convenient return to
hand feeding in case of an emergency. As shown in Fig. 9 there is a
track-way at right angles to the line of furnaces. The car hoisted up
the incline is landed on a transfer carriage, on which, after detaching
the cable, it can be moved over the tops of the furnaces by means of
a tail-rope system. The gage of the charge-car is 4 ft. 9 in.; of
the transfer carriage, 11 ft. 8 in. A switch at the lower end of the
incline permits two charge-cars to be employed, one being filled while
the other is making the trip. In sending down the empty car a hand
winch is necessary to start it from the transfer carriage. Figs. 10 and
11 show the charge-car; Fig. 12 the transfer carriage.
[Illustration: FIG. 8.—Pueblo System. (Sectional diagrams of furnace
top.)]
The charge-car is 10 × 4 × 3.5 ft., and has capacity for 6 tons of ore,
flux, slag and fuel, the total of ore and flux being usually 8800 lb.
Its bottom is flat, consisting of two doors, hinged along the sides
and kept closed by means of chains wound about a longitudinal windlass
on top of the car. The charging pits are decked with iron plates,
leaving a slot along the center of each car exactly like the slot in
the furnace top. The loaded ore-buggies are taken from the wheelers by
two men, who carefully distribute the contents of each buggy along the
whole length of the charge-car by dragging it along the slot while in
the act of dumping. Each buggy contains but one ingredient; they follow
one another in a prescribed order, so as to secure thin layers in the
charge-car. The coke is divided into three or more layers.
[Illustration: FIG. 9.—East Helena System. (Vert-longitudinal section
and plan of incline.)]
[Illustration: FIG. 10.—East Helena Charge-car. (Side elevation.)]
The first few trials of this device were not satisfactory. The furnaces
quickly showed over-fire, and decreased lead output, which would not
yield to any remedy except a return to hand feeding. The total charge
being dropped in the center of the furnace, a central core of fines
was produced, the lumps tending to roll toward the walls. This wrong
tendency was emphasized by the presence of the chains supporting
the bottom of the charge-car. On unwinding them to dump the car,
the doors were prevented from dropping by the wedging of the chains
in the charge, which in turn arched itself more or less against the
sides of the car; hence the doors opened but slowly, and often had to
be assisted by an attendant with a bar. In consequence of this slow
opening, considerable fine ore sifted out first and formed a ridge in
the center of the furnace, from the slopes of which the coarser part of
the charge, the last to fall, naturally rolled toward the sides. This
fact, determined during a visit of the writer in April, 1899, proved
to be the key to the situation. The attendant operating the tail-rope
mechanism was instructed to move the transfer carriage rapidly backward
and forward over the slot while the first one-third or one-half of
the charge was dropping, and during the rest of the discharge to let
the car stand directly over the slot and permit the coarser material
to fall in the center of the furnace. Two piles of comparatively fine
material were thus left on the charge-floor, one on each side of the
slot. These were subsequently fed in by hand, with instructions to
throw the material well to the sides of the furnace.
[Illustration: FIG. 11.—East Helena Charge-car. (Plan.)]
The furnaces were running very hot on top when this modified procedure
was begun. In a few hours the over-fire had disappeared; the lead
output was increasing; and the furnaces were running normally. This was
done about May 1, 1899, and from that time until about February 20,
1900, the Hixon feed, as modified above, was continuously in operation.
In October, 1898, with three furnaces in operation and hand feeding,
the labor cost per furnace was $42.06 per day; in October, 1899, with
the same number of furnaces and mechanical feeding, it was $41 per day,
the saving being only 0.6c. per ton of charge.
[Illustration: FIG. 12.—East Helena Charge-car and Transfer Carriage.
(Elevation.)]
[Illustration: FIG. 13.—East Helena System, with spreader and curtains.
(Experimental form.)]
_Dwight Spreader and Curtain._—In January, 1900, the writer again
had occasion to visit the East Helena plant, to investigate why a
certain cheap local coke could not be used successfully instead of
expensive Eastern coke. Strange as it may seem, the peculiar behavior
of the cokes was traced to improper feeding of the furnaces. Further
study of the mechanical feeding system, then in operation for nine
months, showed that it was far from perfect, and it appeared desirable
to design a spreader which would properly distribute the material
discharged from the Hixon car and dispense with hand feeding entirely.
An experimental construction was arranged, as shown in Fig. 13. The
flanged cast-iron plates around the feeding slot were pushed back and
a roof-shaped spreader, with slopes of 45 deg., was set in the gap,
leaving openings about 8 in. wide on each side. The plan provided for
two iron curtains to be hung, one on each side of the spreader, and so
adjusted that the fine ore sliding down the spreader would clear the
edge of the curtain and shoot toward the sides of the furnace, while
the coarse ore would strike the curtain and rebound toward the center
of the furnace. The classification effected in this manner was capable
of adjustment by raising or lowering the curtain. This arrangement was
found to work surprisingly well. The first furnace equipped with it
immediately showed improvement. It averaged better in speed, with lower
blast, lower lead in slag and matte, and better bullion output than
the other furnaces operating under the old system. The success of the
spreader and curtain being established, the furnaces were provided with
permanent constructions, the only modifications being that the ridge of
the spreader was lowered to correspond with the level of the floor and
the curtains were omitted, the feeding being apparently satisfactory
without their aid. In their absence, the lowering of the spreader was a
proper step, as it distributed the material fully as well, and caused
less abrasion of the walls. The final form is shown approximately
in Fig. 14. It has given complete satisfaction at East Helena since
February, 1900, and has been adopted as the basis for the mechanical
feeding device in the new plant of the American Smelting and Refining
Company at Salt Lake, Utah.
[Illustration: FIG. 14.—East Helena System. (Final form, approximate.)]
_Comparison of Systems._—In mechanical design the Pueblo system
is better than the East Helena, being simpler in construction and
operation. No time is lost in attaching and changing cables, operating
transfer carriage, etc. In both systems the track runs directly over
the tops of the furnaces, and this is an inconvenience when furnace
repairs are under way. The Pueblo car is the simpler, and makes the
round trip in about half the time of a car at East Helena, so the two
cars of the latter do not make much difference in this respect. The
system of filling the charge-car at Pueblo is also the quicker. It may
be estimated roughly that per ton of capacity it takes 2.5 to 3 times
as long to fill the East Helena car; and this means longer waiting on
the part of the wheelers, and consequently greater cost of moving the
material, representing probably 7 or 8c., in favor of Pueblo, per ton
of charge handled. However, both systems are wasteful of labor. As to
furnace results, it is believed that the better distribution of the
charge in the East Helena system leads to greatly increased regularity
of furnace running, less tendency to over-fire, some economy in fuel,
less accretions on the furnace walls and larger metal savings. If the
half of these conclusions are true, the difference of 7 or 8c. per ton
in favor of the Pueblo system, which can be traced almost entirely
to the cost of filling the charge-car, sinks into insignificance
in comparison with the important advantages of having the furnaces
uniformly and correctly fed.
_True Function of the Charge-Car._—The radically essential feature of a
mechanical feeding device is that part which automatically distributes
the material in the furnace, whatever approximate means may have been
used to effect the delivery.
Taking a hasty review of the numerous feeding devices that have
been tried in lead-smelting practice, we cannot but remark the fact
that those which depended upon dumping the charge into the furnace
from small buggies or barrows failed generally to secure a proper
classification and distribution of coarse and fine, and, consequently,
were abandoned as unsuccessful, while the adoption of the idea of the
charge-car for transporting the material to the furnace in large units
seems to have been coincident with a successful outcome. It is natural
enough, therefore, that the car should be regarded by many as the vital
feature. This view of the question is not, however, in accordance
with the true perspective of the facts, and merely limits the field
of application in an entirely unnecessary way. It must be apparent
that the essential function of the charge-car is cheap and convenient
transportation. The distribution of the charge is an entirely different
matter, in which, however, the charge-car may be made to assist, as
in the Pueblo system; or entirely distinct and special means may be
employed for the distribution, as in the East Helena system.
To follow the argument to its conclusion, let us imagine for the moment
that the East Helena plant were arranged on the terrace system, with
the furnace tops on a level with the floor of the ore-bins. Certain
precautions being observed, the spreader would give as good results
with small units of charge delivered by buggies as it now does with the
large units delivered by the charge-car, and the expense of delivery
to the furnaces would be practically no more than it now is to the
charge-car pit. The furnace top would, of course, have to be arranged
so that the buggies, in discharging, could be drawn along the slot,
so as to give the necessary longitudinal distribution parallel to the
furnace walls, just as is now done in filling the charge-car. The ends
of the spreader, if built like a hipped roof, would secure proper
feeding of the front and back.
Thus, by eliminating the charge-car, and with it the necessity for
powerful hoisting machinery, with its expensive repairs and operating
costs, we may greatly simplify the problem of mechanical feeding, and
open the way for the adoption of successful automatic feeding in many
existing plants where it is now considered impracticable.
COST OF SMELTING AND REFINING
BY MALVERN W. ILES
(August 18, 1900)
In the technical literature of lead smelting there is a lamentable lack
of data on the subject of costs. The majority of writers consider that
they have fulfilled their duties if they discuss in full detail the
chemical and engineering sides of the subject, leaving the industrial
consideration of cost to be wrought out by experience. When an engineer
or metallurgist collects data on the costs involved in the various
smelting operations, he generally hesitates to give this special
information to the public, as he regards it as private, or reserves it
as stock in trade to be held for his own use.
The following tables of cost have been compiled from actual results
of smelting and refining at the Globe works, Denver, Colo., and are
offered in the hope that they will prove a valuable addition to the
literature of lead smelting. These results are offered tentatively,
and, while true for the periods stated, they require considerable
adjustment to meet the smelting conditions of the present time.
COST OF HAND-ROASTING PER TON (2000 LB.) OF ORE
1887 $3.975
1888 4.280
1889 4.120
1890 3.531
1891 3.530
1892
1893
1894 3.429
1895 2.806
1896 2.840
1897 2.740
1898 2.620
At first the roasting was done mainly by hand roasters; later two
Brown-O’Harra mechanical furnaces were used, and the cost was reduced,
but not to the extent usually conceded to this type of furnace, as the
large amount of repairs and the consequent loss of time diminished
the apparent gain due to greater output. The figures quoted above may
be considered somewhat higher than the average, as the roasters were
charged in proportion with expenses of general management, office, etc.
In viewing the yearly reduction of costs one must take into
consideration many changes in the furnace construction and working, as
well as the items of labor, fuel, etc. From 1887 to 1899 the principal
changes in the construction of the hand-roasting furnaces consisted
in an increase of width, 2 ft., which allowed an addition of 200 lb.
to each ore charge, and corresponded to a total increase per furnace
of 1200 lb. in 24 hours. In the working of the charge an important
change was made in the condition of the product. Formerly the material
was fused in the fusion-box and drawn from the furnace in a fused or
slagged condition; and while this gave an excellent material for the
subsequent treatment in the shaft furnace in that there was very little
dusting of the charge, and a considerable increase in the output of the
furnace, the disadvantages of large losses of lead and silver greatly
over-balanced the advantages, and called for an entire abandonment of
the fusion-box. As a result of experience it was found that the best
condition of product is a semi-fused or sintered state, in which the
particles of roasted ore have been compressed by pounding the material,
which has been drawn into the slag pots, with a heavy iron disk. The
amount of “fines” under these conditions is quite small and depends
upon the percentage of lead in the ore, the degree of heat employed,
and the extent of the compression.
The total cost was partly reduced from the lessened labor cost
following the financial disturbance of 1893, and partly from the
reduction in the fuel cost, the former expensive lump coal being
replaced by the slack coals from southern Colorado.
The comparison of the cost of labor by the two methods shows a gain of
54c. a ton in favor of the mechanical furnaces. However, I consider
that this gain is a costly one, and is more than offset by the large
amount of high-grade fuel required, and the expense of repairs not
shown in the following table. Indeed, I believe that at the end of five
or ten years the average cost of roasting per ton by the hand roasters
will be even smaller than by these mechanical roasters.
To illustrate the details of roasting cost and to furnish a comparison
of the hand roasters and mechanical furnaces, the following table has
been prepared:
DETAILS OF AVERAGE MONTHLY COST FOR 1898 OF HAND ROASTERS AND
MECHANICAL FURNACES
───────────┬────────┬───────┬─────────────────────┬─────────────────────
│ │ │ HAND ROASTERS │ BROWN-O’HARRA
│ │ │ │ MECHANICAL FURNACES
│ TOTAL │ TONS ├──────┬──────┬───────┼──────┬──────┬───────
Month │ TONS │ROASTED│LABOR │ COAL │GENERAL│LABOR │COAL │GENERAL
│ROASTED │PER DAY│ $ │ $ │EXPENSE│ $ │ $ │EXPENSE
│ │ │ │ │ $ │ │ │ $
───────────┼────────┼───────┼──────┼──────┼───────┼──────┼──────┼───────
January │ 5,691 │ 184 │ 1.47 │ 0.53 │ 0.80 │ 0.92 │ 0.80 │ 1.32
February │ 5,677 │ 203 │ 1.44 │ 0.44 │ 0.99 │ 0.72 │ 0.58 │ 1.01
March │ 5,821 │ 188 │ 1.51 │ 0.53 │ 0.64 │ 0.76 │ 0.64 │ 0.62
April │ 5,472 │ 182 │ 1.47 │ 0.47 │ 0.71 │ 0.80 │ 0.69 │ 0.87
May │ 5,444 │ 176 │ 1.55 │ 0.51 │ 0.84 │ 0.80 │ 0.69 │ 0.81
June │ 4,859 │ 162 │ 1.58 │ 0.48 │ 0.71 │ 0.90 │ 0.68 │ 1.17
July │ 5,691 │ 184 │ 1.59 │ 0.48 │ 0.75 │ 0.72 │ 0.56 │ 0.64
August │ 5,910 │ 191 │ 1.55 │ 0.46 │ 0.83 │ 0.72 │ 0.55 │ 0.75
September │ 5,677 │ 189 │ 1.55 │ 0.45 │ 0.74 │ 0.73 │ 0.55 │ 0.67
October │ 6,254 │ 202 │ 1.48 │ 0.49 │ 0.72 │ 0.65 │ 0.50 │ 0.60
November │ 6,291 │ 213 │ 1.42 │ 0.47 │ 0.80 │ 0.66 │ 0.53 │ 0.70
December │ 5,874 │ 198 │ 1.45 │ 0.48 │ 0.78 │ 0.79 │ 0.63 │ 0.81
├────────┼───────┼──────┼──────┼───────┼──────┼──────┼───────
Average │ │ │ 1.50 │ 0.48 │ 0.77 │ 0.76 │ 0.62 │ 0.83
Total │ │ │ │ │ 2.75 │ │ │ 2.21
───────────┴────────┴───────┴──────┴──────┴───────┴──────┴──────┴───────
_Cost of Smelting._—The lead-ore mixtures of the United States, in
addition to lead, contain gold, silver and generally copper, and are
treated to save these metals. The total cost of smelting is made up of
a large number of items. The questions of locality and transportation,
fuel, fluxes and labor are the principal factors, to which must be
added the handling of the material to and from the furnace; the
furnace itself, its size, shape, and method of smelting, the volume
and pressure of blast, etc. The following table of costs, from 1887 to
1898, shows in a general way the great advance that has been made in
the development of smelting, and the consequent reduction in cost per
ton of ore treated:
AVERAGE COST OF SMELTING, PER TON
1887 $4.644
1888 4.530
1889 4.480
1890 4.374
1891 4.170
1892 4.906
1893 3.375
1894 3.029
1895 2.786
1896 2.750
1897 2.520
1898 2.260
In connection with this table of smelting cost should be considered the
changes developed during the interval 1887-1889, outlined as follows:
CONDITIONS OF SMELTING IN 1886 AND 1899 CONTRASTED TO SHOW THE PROGRESS
OF DEVELOPMENT
────┬───────────────┬────────────┬───────────────┬─────────────┐
│AREA OF FURNACE│ HEIGHT OF │BLAST PRESSURE,│ FORE HEARTH │
│ AT TUYERES, │CHARGE FROM │ LB. PER │CAPACITY, CU.│
│ IN. │TUYERES, FT.│ SQ. IN. │ FT. │
────┼───────────────┼────────────┼───────────────┼─────────────┤
1886│ 30 × 100 │ 11 │ 1 │ 6 │
│ │ │ │ │
│ │ │ │ │
1899│ 42 × 140 │ 16 │ 3 to 4 │ 128 │
│ │ │ │ │
────┴───────────────┴────────────┴───────────────┴─────────────┘
────┬────────────┬────────┬───────────────┬───────────────┐
│ SLAG │ FUEL │ SLAG REMOVED, │MATTE REMOVED, │
│ SETTLED │ │ LB. PER TRIP │ LB. PER │
────┼────────────┼────────┼───────────────┼───────────────┤
1886│ │ │ │ TRIP │
│ In pots │Charcoal│ By hand │ By hand │
│ │ │ 280 │ 200 │
1899│ │ │ │ │
│In furnaces │ Coke │ By locomotive │ By horse │
│ │ │ 3000-6000 │ 2000-3000 │
────┴────────────┴────────┴───────────────┴───────────────┘
I believe that there is room for further improvement in the
substitution of mechanical transportation within the works for hand
labor, and that the fuel cost can be materially reduced by replacing
the coke, which at present contains 16 to 22 per cent. of ash, by a
fuel of purer and better quality.
_Cost of Refining by the Parkes Process._—In general it may be stated
that the average cost of refining base bullion is from $3 to $5 a
ton. This amount is based on the cost of labor, spelter, coal, coke,
supplies, repairs and general expenses. When the additional items
of interest, expressage, brokerage and treatment of by-products are
considered, which go to make up the total refining cost, the amount may
be stated approximately as $10 per ton of bullion treated.
Variations in the cost occur from time to time, and are due to several
causes, principally the irregularity of the bullion supply and its
consequent effect on the work of the plant. When the amount of bullion
available for treatment is small, the plant cannot be run to its
maximum capacity, and the cost per ton will naturally be increased. To
illustrate this variation, the average cost per ton of base bullion
refined during nine months in 1893 was:
January, $4.864; February, $5.789; March, $5.024; April, $3.915; May,
$5.094; June, $4.168; July, $4.231; August, $4.216; September, $5.299.
The yearly variation shows but little change, as the average cost per
ton was for 1893, $4.75; for 1894, $3.99; for 1895, $4.21; for 1896,
$3.90. In considering the total cost of refining, the additional
factors of interest, expressage, parting, brokerage, and reworking of
by-products must be considered. As the doré silver is treated at the
works or elsewhere, so will the total cost be less or greater. The
following table gives the cost in detail, when the parting is done at
the same works:
AVERAGE MONTHLY COST OF REFINING PER TON OF BULLION TREATED
─────────────────────┬────────────┬────────────┬────────────┬─────────
ITEMS │ 1895 │ 1895 │ 1896 │ AVERAGE
│JAN. TO JULY│JULY TO DEC.│JAN. TO JULY│
─────────────────────┼────────────┼────────────┼────────────┼─────────
Labor │ $2.351 │ $1.718 │ $1.836 │ $1.968
Spelter │ 0.757 │ 0.840 │ 0.987 │ 0.861
Coal │ 0.585 │ 0.442 │ 0.461 │ 0.496
Coke │ 0.634 │ 0.418 │ 0.511 │ 0.521
Supplies, repairs and│ │ │ │
general expenses │ 0.343 │ 0.273 │ 0.252 │ 0.289
Interest │ 1.808 │ 1.075 │ 1.070 │ 1.317
Expressage │ 1.360 │ 1.015 │ 0.882 │ 1.085
Parting and brokerage│ 2.483 │ 2.084 │ 1.796 │ 2.121
Reworking by-products│ 1.567 │ 1.286 │ 1.625 │ 1.492
├────────────┼────────────┼────────────┼─────────
Totals │ $11.888 │ $9.151 │ $9.420 │ $10.151
Tons bullion refined │5,511.58 │9,249.07 │10,103.43 │8,287.99
─────────────────────┴────────────┴────────────┴────────────┴─────────
An analysis of the different items of cost is important, and a brief
summary is given below.
_Labor and Attendance._—The cost for this item varies but little from
year to year, and its reduction depends, for the most part, on a larger
yield per man rather than on a reduction of wages. If a man at the same
or slightly increased cost can give a larger output, so will the labor
cost per ton be diminished. This result is accomplished by enlarging
the furnace capacity and by using appliances which will handle the
bullion and its products in an easier and quicker manner. The small
size of the furnaces, settlers and retorts used at modern refineries is
open to criticism; I believe that great improvement can be made in this
direction.
_Spelter._—The cost of this item varies with the market conditions,
and will probably be changed but little in the future, as the amount
necessary per ton of bullion seems to be fixed.
_Coal._—The amount required per ton of bullion is fairly constant, and
while lessened cost for fuel may be attained by the substitution of oil
or gaseous fuel, the fuel cost in comparison with the aggregate cost is
very small, and leaves little opportunity for improvement in this line.
_Supplies._—This item includes brooms, shovels, wheelbarrows, etc., and
the amount is small and fairly constant from year to year.
_Repairs._—This item is quite small in works properly constructed;
and in this connection I wish to call particular attention to the
floor covering, which should be made of cast-iron plates from 1.5 to
2 in. thick, and placed on a 2 to 3 in. layer of sand spread over the
well-tamped and leveled ground. The constant patching of brick floors
is not only an annoyance, but is costly from the additional labor
required. Furthermore, a brick floor does not permit a close saving of
the metallic scrap material.
It will be found economical in the long run to protect all exposed
brickwork of furnaces or kettles with sheet iron.
In the construction of the refinery building I should advise brick
walls except at the end or side, where there is the greatest likelihood
of future extension; here corrugated iron may be used. The roof should
not be made of corrugated iron, as condensed or leakage water is liable
to collect and drop on those places where water should be scrupulously
avoided. The presence of water in a mold at the time of casting, even
though small in amount, will cause explosions and will scatter the
molten lead, endangering the workmen.
The item of repair for the ordinary corrugated iron roof may be
diminished by constructing it of 1 in. boards with intervening spaces
of half an inch, the whole overlaid with tarred felt, and covered with
sheets of iron at least No. 27 B. W. G., painted with graphite paint
and joined together with parallel rows of ribbed crimped iron.
_General Expenses._—This item is generally constant, and calls for no
special comment.
_Interest._—This important item is, as a rule, considerable, as the
stock of bullion and other gold-and silver-bearing material is quite
large. For this reason special attention should be given to prevent
the accumulation of stock or by-products. The occasional necessity of
additional capital to run the business should preferably be met by an
increase of working capital, rather than by a direct loan.
_Expressage._—This item, as a rule, is large, and should be taken into
consideration in the original plans for the location of the refining
works.
_Parting._—The item of parting and brokerage is the largest of the
refinery costs, and for obvious reasons a modern smelting plant should
have a parting plant under its own control.
_The Working of the By-Products._—This constitutes a large item of
cost, and considerable attention should be devoted to the improvement
of present methods, which seem faulty, slow and expensive.
_Summary._—The items of smaller cost with their respective amounts per
ton of base bullion treated are: Spelter, $0.85; coal, $0.50; coke,
$0.50; supplies, repairs and general expenses, $0.35; total, $2.10. It
is doubtful whether much improvement can be made in the reduction of
these costs.
The items of larger cost are: Labor, $2; interest, $1.32; expressage,
$1.10; parting and brokerage, $2; reworking by-products, $1.50; total,
$7.92. The general manager usually attends to the items of interest,
expressage and brokerage, leaving the questions of labor and working of
by-products to the metallurgist.
The cost quoted for smelting practice, as employed at Denver, will
differ necessarily from those at other localities, where the cost of
labor, freight rates on spelter, fuel, etc., are changed. Refining
can doubtless be done at a lower cost at points along the Mississippi
River, and even more so at cities on the Atlantic seaboard, as Newark
or Perth Amboy, N. J.
The consolidation of many of the more important smelting plants of the
United States under one management will doubtless alter the figures of
cost given above, particularly as the interest cost there stated is at
the high rate of 10 per cent., a condition of affairs now changed to 5
per cent. Other factors have lessened the cost of refining; the bullion
produced at the present time is softer, or contains a smaller amount
of impurities, and admits of easier working with shorter time and less
labor. By proper management larger tonnages are turned out per man, and
the Howard stirrer and Howard press have simplified and cheapened the
working of the zinc skimmings. To illustrate the comparatively recent
conditions of cost I have compiled the following table for each month
of the year 1898:
COST OF REFINING DURING 1898, INCLUDING LABOR, SPELTER, COAL, COKE,
SUPPLIES, REPAIRS AND GENERAL EXPENSES.
January $3.59
February 3.28
March 3.26
April 3.59
May 3.38
June 3.56
July 3.65
August 3.54
September 3.35
October 3.45
November 3.20
December 3.56
Average cost during the year, $3.45.
It is understood, of course, that these figures do not include cost of
interest, expressage, parting, brokerage and reworking of by-products.
[Although this article refers to conditions in 1898, since which time
there have been improvements in practice, the latter have not been of
radical character and the figures given are fairly representative of
present conditions.—EDITOR.]
SMELTING ZINC RETORT RESIDUES[13]
BY E. M. JOHNSON
(March 22, 1906)
The following notes were taken from work done at the Cherokee
Lanyon Smelter Company, Gas, Kansas, in 1903. It was practically an
experiment. The furnace was only 36 × 90 in. at the crucible, with a
10 in. side bosh and a 6 in. end bosh. There were five tuyeres on each
side with a 3 in. opening. The side jackets measured 4.5 ft. × 18 in.
The distance from top of crucible to center of tuyeres was 11.5 in.
The blast was furnished by one No. 4½ Connellsville blower. The
furnace originally was only 11 ft. from the center of tuyeres to the
feed-floor, and had only been saving about 60 per cent. of the lead.
This loss of lead, however, was not entirely due to the low furnace.
As no provision had been made to separate the slag and matte, upon
assuming charge I raised the feed-floor 3 ft., thereby changing the
distance from the tuyere to top of furnace from 11 ft. to 14 ft. Matte
settlers were also installed. These two changes raised the percentage
of lead saved to 92, as shown by monthly statements. The furnace being
small, and a high percentage of zinc oxide on the charge, the campaigns
were naturally short. The longest run was about six weeks. This was
made on some residue that had been screened from the coarse coal, and
coke, and had weathered for several months. This particular residue
also carried about 10 per cent. lead. The more recent residue that had
not been screened and weathered, and was low in lead, did not work so
well. Although these residues consisted of a large proportion of coal
and coke, it seemed impossible to reduce the percentage of good lump
coke on the charge lower than 12.5 or 13 per cent. At the same time the
reducing power of the residue was strong, and with the normal amount of
coke caused some trouble in the crucible.
When residue containing semi-anthracite coal was smelted, the saving
in lead dropped, and the fire went to the top of the furnace, burning
with a blue flame, thereby necessitating the reduction of this class of
material. This residue had been screened through a five-mesh screen,
and wet down in layers, becoming so hard that it had to be blasted.
The low saving of lead with this class of material was a surprise,
as it has been claimed that the substitution of part of the fuel by
anthracite coal did not affect the metallurgical operations of the
furnace.
The slag was quite liquid and flowed very well at all times. However,
there was a marked variation in the amount at different tappings. This,
I am satisfied, was not due to irregular work on the furnace, but may
be accounted for in the following manner. The residue (not screened or
weathered to any extent), consisting approximately of one-half coal
and coke, was very bulky, and while there was about 35 per cent. of it
on the charge by weight, there was over 50 per cent. of it by bulk,
not including slag and coke. In feeding, therefore, it was a difficult
matter to mix the whole of it with the charge. Several different ways
of feeding the furnace were tried. The one giving the most satisfactory
results was to feed nearly all of the residue along the center of the
furnace, in connection with the lime-rock, coarse ore and coarse iron
ore, and the fine and easy smelting ores along the sides. The slag was
spread uniformly over the whole furnace, while the sides were favored
with the coke. The charge would drop several inches at a time, going
down a little faster in the center than on the sides.
It is possible that a small proportion of the residue in connection
with the easy smelting, leady, neutral ore, iron ore and lime-rock
formed the type of slag marked No. 1.
───┬───────┬──────┬─────┬──────┬─────┬─────┬────
│ SiO₂ │ FeO │ MnO │ CaO │ ZnO │ Pb │ Ag
───┼───────┼──────┼─────┼──────┼─────┼─────┼────
1 │ 33.7 │ 34.1 │ 1.0 │ 16.5 │ 7.5 │ 0.9 │ 0.7
2 │ 31.0 │ 36.1 │ 1.2 │ 16.0 │ 9.6 │ 1.3 │
───┴───────┴──────┴─────┴──────┴─────┴─────┴────
This being tapped with a good flow of slag, the charge would drop,
bringing a proportionately large amount of residue in the fusion zone
which formed the type of slag marked No. 2. There was also a marked
variation in the slag-shells from different pots. The above cited
irregularities of course exist to a certain extent in any blast furnace.
AVERAGE ANALYSIS OF MATERIALS SMELTED
NAME ROW NAME ROW
Mo. iron ore A Roasted matte[15] F
Lime rock B Barrings G
Mo. galena C Coke ash H
Av. of beds D Coke[16] J
Residue[14] E
────┬──────┬─────┬────┬────┬────┬─────┬─────┬────┬────┬───┬────┬────
│ SiO₂ │ FeO │CaO │MgO │ZnO │Al₂O₃ │Fe₂O₃ │ S │ Pb │Cu │ Ag │ Au
────┼──────┼─────┼────┼────┼────┼─────┼─────┼────┼────┼───┼────┼────
A │ 10.0 │ 65.0│ │ │ │ │ │ │ │ │ │
B │ 1.5 │ │52.0│ │ │ │ │ │ │ │ │
C │ 1.5 │ 2.4│ │ │ 9.5│ │ │11.0│74.0│ │ │
D │ 50.8 │ 16.2│ │ │ 4.6│ │ │ 3.3│ 9.1│ │ │
E │ 10.5 │ 38.5│ │ │18.0│ │ │ 4.8│ 2.2│1.0│10.0│0.03
F │ 9.0 │ 48.0│ 3.0│ │10.0│ │ │ 4.0│ 9.9│3.0│21.0│0.06
G │ 18.8 │ 24.4│ 5.0│ │14.5│ │ │ 6.0│25.4│ │13.0│0.07
H │ 27.0 │ │14.9│ 4.5│ │ 19.7│ 31.6│ │ │ │ │
│ H₂O │ V.M.│F.C.│ Ash│ S │ │ │ │ │ │ │
J │ 1.2 │ 2.3 │85.7│11.1│ 0.9│ │ │ │ │ │ │
─────┴──────┴─────┴────┴────┴────┴─────┴─────┴────┴────┴───┴────┴────
ANALYSIS OF BULLION, SLAG AND MATTE PRODUCED
│/-BULLION-\ /—————————————SLAG———————————————-\/————-MATTE————-\
│ Ag │ Au │SiO₂ │FeO │MnO│CaO │ZnO │ Pb │ Ag │ Ag │ Au │Pb │Cu
├──────┼────┼─────┼────┼───┼────┼────┼────┼────┼────┼────┼───┼───
Feb. │ 90.0 │1.15│31.2 │35.9│1.0│14.5│10.3│0.88│0.98│19.0│0.04│8.7│1.5
March│ 93.1 │1.63│31.3 │37.2│1.0│13.9│11.1│0.71│1.30│21.0│0.06│8.0│2.5
April│104.3 │1.59│29.8 │37.7│2.7│13.9│11.4│0.52│1.40│23.0│0.07│7.0│3.5
May │ 90.0 │1.24│30.0 │37.3│2.2│14.1│ 9.3│0.86│1.10│25.4│0.07│5.1│4.0
July │ 78.7 │1.00│32.2 │37.4│1.0│13.9│ 9.8│0.50│1.15│21.3│0.03│8.9│4.0
Aug. │ 90.8 │1.21│31.2 │37.1 1.7│13.7│ 9.6│1.10│1.60│23.1│0.08│9.8│3.0
Sept.│ 65.3 │2.58│32.0 │39.7│0.8│14.1│ 8.1│0.80│1.30│18.6│0.06│7.6│2.3
├──────┼────┼─────┼────┼───┼────┼────┼────┼────┼────┼────┼───┼───
Avge.│ 87.5 │1.49│31.1 │37.5│1.5│14.1│10.0│0.77│1.26│21.6│0.06│7.8│3.0
─────┴──────┴────┴─────┴────┴───┴────┴────┴────┴────┴────┴────┴───┴───
MONTHLY RECORD OF FURNACE OPERATIONS
─────────┬──────┬───────┬─────────┬─────────┬─────────┬─────────┐
│BLAST │ TONS │PER CENT.│PER CENT.│PER CENT.│PER CENT.│
│OUNCES│ PER │ PB. ON │ COKE ON │ SLAG ON │ S ON │
│ │ F.D. │ CHARGE │ CHARGE │ CHARGE │ CHARGE │
─────────┼──────┼───────┼─────────┼─────────┼─────────┼─────────┤
Feb. │ 21 │ 42.5 │ 9.0 │ 12.0 │ 30.0 │ 3.7 │
March │ 21 │ 44.8 │ 9.7 │ 13.5 │ 37.0 │ 4.0 │
April │ 21 │ 43.7 │ 9.0 │ 13.5 │ 35.0 │ 4.3 │
May │ 21 │ 49.4 │ 10.0 │ 13.5 │ 30.0 │ 3.5 │
July │ 17 │ 41.0 │ 9.8 │ 12.5 │ 34.0 │ 3.8 │
August │ 18 │ 47.0 │ 9.3 │ 13.0 │ 32.0 │ 3.7 │
Sept.[17]│ 15 │ 51.0 │ 7.3 │ 13.0 │ 30.0 │ 2.8 │
─────────┼──────┼───────┼─────────┼─────────┼─────────┼─────────┤
Average │ │ 45.6 │ 9.1 │ 13.0 │ 32.6 │ 3.7 │
─────────┴──────┴───────┴─────────┴─────────┴─────────┴─────────┘
────────┬────────┬─────────────────────┐
│ MATTE │ SAVING │
│PRODUCED│ AG AU PB │
────────┼────────┼──────┬───────┬──────┤
Feb. │ 8.0} │ 84.4 │ 83.0 │ 90.3 │
March │ 9.0} │ │ │ │
April │ 10.0 │ 97.9 │ 70.5 │ 96.6 │
May │ 6.5 │ 95.6 │ 109.5 │ 88.8 │
July │ 6.0 │ 97.9 │ 90.0 │ 92.9 │
August │ 6.3 │ 86.2 │ 107.5 │ 87.6 │
Sept. │ 4.6 │ 92.9 │ 94.0 │ 95.6 │
────────┼────────┼──────┼───────┼──────┤
Average │ 7.2 │ 90.8 │ 92.4 │ 92.0 │
────────┴────────┴──────┴───────┴──────┘
I believe that, in smelting residues high in zinc oxide, better
metallurgical results would be obtained by using a dry silicious ore in
connection with a high-grade galena ore, provided the residue be low in
sulphur. This was confirmed to a certain degree in actual practice, as
the furnace worked very well upon increasing the percentage of Cripple
Creek ore on the charge. This would also seem to indicate that alumina
had no bad effect on a zinky slag.
ZINC OXIDE IN SLAGS
BY W. MAYNARD HUTCHINGS
(December 24, 1903)
From time to time, in various articles and letters on metallurgical
subjects in the _Engineering and Mining Journal_, the question of the
removal of zinc oxide in slags is referred to, and the question is
raised as to the form in which it is contained in the slags.
I gather that opinion is divided as to whether zinc oxide enters into
the slags as a combined silicate, or whether it is simply carried into
them in a state of mechanical mixture.
For many years I have taken great interest in the composition of slags,
and have studied them microscopically and chemically. The conclusion to
which I have been led as regards zinc oxide is, that in a not too basic
slag it is originally mainly, if not wholly, taken up as silicate along
with the other bases. On one occasion, one of my furnaces for several
days produced a slag in which beautiful crystals of willemite were
very abundant, both free in cavities and also imbedded throughout the
mass of solid slag, as shown in thin sections under the microscope. In
the same slag was a large amount of magnetite, all of which contained
a considerable proportion of zinc oxide combined with it. Magnetite
crystals, separated out from the slag and treated with strong acid,
yielded shells of material retaining the form of the original mineral,
rich in zinc oxide; an inter-crystallized zinc-iron spinel, in fact.
I have seen and separated zinc-iron spinels very rich in zinc oxide
from other slags. They have been seen in the slags at Freiberg; and
of course everybody knows the very interesting paper by Stelzner and
Schulze, in which they described the beautiful formations of spinels
and willemite in the walls of the retorts of zinc works.
I think there is thus good ground for concluding that zinc oxide is
slagged off as combined silicate, and that free oxide does not exist
in slags; though zinc oxide does occur in them after solidification,
combined with other oxides, in forms ranging from a zinkiferous
magnetite to a more or less impure zinc-iron, or zinc-iron-alumina
spinel, these minerals having crystallized out in the earlier stages of
cooling.
The microscope showed that the crystals of willemite, mentioned above,
were the first things to crystallize out from the molten slag. The main
constituent was well-crystallized iron-olivine-fayalite.
PART V
LIME-ROASTING OF GALENA
THE HUNTINGTON-HEBERLEIN PROCESS
(July 6, 1905)
It is a fact, not generally known, that the American Smelting and
Refining Company is now preparing to introduce the Huntington-Heberlein
process in all its plants, this action being the outcome of extensive
experimentation with the process. It is contemplated to employ the
process not only for the desulphurization of all classes of lead ore,
but also of mattes. This is a tardy recognition of the value of a
process which has been before the metallurgical profession for nine
years, the British patent having been issued under date of April
16, 1896, and has already attained important use in several foreign
countries; but it will be the grandest application in point of
magnitude.
The Huntington-Heberlein is the first of a new series of processes
which effect the desulphurization of galena on an entirely new
principle and at great advantage over the old method of roasting.
They act at a comparatively low temperature, so that the loss of lead
and silver is reduced to insignificant proportion; they eliminate the
sulphur to a greater degree; and they deliver the ore in the form of
a cinder, which greatly increases the smelting speed of the blast
furnace. They constitute one of the most important advances in the
metallurgy of lead. The roasting process has been the one in which
least progress has been made, and it has remained a costly and wasteful
step in the treatment of sulphide ores. In reducing upward of 2,500,000
tons of ore per annum, the American Smelting and Refining Company is
obliged to roast upward of 1,000,000 tons of ore and matte.
The Huntington-Heberlein process was invented and first applied at
Pertusola, Italy. It has since been introduced in Germany, Spain, Great
Britain, Mexico, British Columbia, Tasmania, and Australia, in the last
at the Port Pirie works of the Broken Hill Proprietary Company. Efforts
were made to introduce it in the United States at least five years ago,
without success and with little encouragement. The only share in this
metallurgical improvement that this country can claim is that Thomas
Huntington, one of the inventors, is an American citizen, Ferdinand
Heberlein, the other, being a German.
LIME-ROASTING OF GALENA
(September 22, 1905)
The article of Professor Borchers (see p. 116) is, we believe, the
first critical discussion of the reactions involved in the new methods
of desulphurizing galena, as exemplified in the processes of Huntington
and Heberlein, Savelsberg, and Carmichael and Bradford, although
the subject has been touched upon by Donald Clark, writing in the
_Engineering and Mining Journal_. It is perfectly obvious from a study
of the metallurgy of these processes that they introduce an entirely
new principle in the oxidation of galena, as Professor Borchers points
out. Inasmuch as there are already three of these processes and are
likely to be more, it will be necessary to have a type-name for this
new branch of lead metallurgy. We venture to suggest that it may be
referred to as the “lime-roasting of galena,” inasmuch as lime is
evidently a requisite in the process; or, at all events, it is the
agent which will be commonly employed.
When the Huntington-Heberlein process was first described, it did not
even appear a simplification of the ordinary roasting process, but
rather a complication of it. The process attracted comparatively little
attention, and was indeed regarded somewhat with suspicion. This was
largely due to the policy of the company which acquired the patent
rights in refusing to publish the technical information concerning it
that the metallurgical profession expected and needed. The history of
this exploitation is another example of the disadvantage of secrecy
in such matters. The Huntington-Heberlein process has only become
thoroughly established as a new and valuable departure in metallurgy, a
departure which is indeed revolutionary, nine years after the date of
the original patent. In proprietary processes time is a particularly
valuable element, inasmuch as the life of a patent is limited.
From the outset the explanation of Huntington and Heberlein as to the
reactions involved in their process was unsatisfactory. Professor
Borchers points out clearly that their conception of the formation of
calcium peroxide was erroneous, and indicates strongly the probability
that the active agent is calcium plumbate. It is very much to be
regretted that he did not go further with his experiments on this
subject, and it is to be hoped that they will be taken up by the
professors of metallurgy in other metallurgical schools. The formation
of calcium plumbate in the process was clearly forecasted, however, by
Carmichael and Bradford in their first patent specification; indeed,
they considered that the sintered product consisted largely of calcium
plumbate.
Even yet, we have only a vague idea of the reactions that occur in
these processes. There is undoubtedly a formation of calcium sulphate,
as pointed out by Borchers and Savelsberg; but that compound is
eventually decomposed, since it is one of the advantages of the
lime-roasting that the sintered product is comparatively low in
sulphur. Is it true, however, that the calcium eventually becomes
silicate? If so, under what conditions is calcium silicate formed? The
temperature maintained throughout the process is low, considerably
lower than that required for the formation of any calcium silicate by
fusion.
Moreover, it is not only galena which is decomposed by the new
method, but also blende, pyrite and copper sulphides. The process is
employed very successfully in the treatment of Broken Hill ore that is
rather high in zinc sulphide, and it is also to be employed for the
desulphurization of mattes. What are the reactions that affect the
desulphurization of the sulphides other than lead?
There is a wide field for experimental metallurgy in connection with
these new processes. The important practical development is that they
do actually effect a great economy in the reduction of lead sulphide
ores.
THE NEW METHODS OF DESULPHURIZING GALENA[18]
BY W. BORCHERS
(September 2, 1905)
An important revolution in the methods of smelting lead ore, which had
to a large extent remained for centuries unchanged in their essentials,
was wrought by the invention of Huntington and Heberlein in 1896. More
especially is this true of the roast-reduction method of treating
galena, which consists of oxidizing roasting in a reverberatory furnace
and subsequent smelting of the roasted product in a shaft furnace.
The first stage of the roast-reduction process, as carried out
according to the old method, viz., the oxidizing roast of the galena,
serves to convert the lead sulphide into lead oxide:
PbS + 3O = PbO + SO₂.
Owing to the basic character of the lead oxide, the production of a
considerable quantity of lead sulphate was of course unavoidable:
PbO + SO₂ + O = PbSO₄.
As this lead sulphate is converted back into sulphide in the
blast-furnace operation, and so adds to the formation of matte, it
has always been the aim (in working up ores containing little or no
copper to be concentrated in the matte) to eliminate the sulphate as
completely as possible, by bringing the charge, especially toward the
end of the roasting operation, into a zone of the furnace wherein
the temperature is sufficiently high to effect decomposition of the
sulphate by silica:
PbSO₄ + SiO₂ = PbSiO₃ + SO₃.
But in the usual mode of carrying out the roast in reverberatory
furnaces, the roasting itself on the one hand, and the decomposition of
the sulphates on the other, were effected only incompletely and with
widely varying results.
Little attention has been paid in connection with the roast-reduction
process to the reaction between sulphates and undecomposed sulphides,
which plays so important a part in the roast-reaction method of lead
smelting. As is well known, lead sulphate reacts with lead sulphide in
varying quantities, forming either metallic lead or lead oxide, or a
mixture of both. A small quantity of lead sulphate reacting with lead
sulphide yields under certain conditions only lead:
PbSO₄ + PbS = Pb₂ + 2SO₂.
Within certain temperature limits this reaction even proceeds with
liberation of heat. In order to encourage it, it is necessary to create
favorable conditions for the formation of considerable quantities
of sulphate right at the beginning of the operation. This was first
achieved by Huntington and Heberlein, but not in the simplest nor in
the most efficient manner. And, indeed, the inventors were not by any
means on the right track as to the character of their process, so far
as the chemical reactions involved are concerned.
At first sight the Huntington-Heberlein process does not even appear
as a simplification, but rather as a complication, of the roasting
operation. For in place of the roast carried out in one apparatus
and continuously, there are two roasts which have to be carried out
separately and in two different forms of apparatus; nevertheless, the
ultimate results were so favorable that the whole process is presumably
acknowledged, without reservation, by all smelters as one of the most
important advances in lead smelting.
It is useful to examine in the light of the German patent specification
(No. 95,601 of Feb. 28, 1897) what were the ideas of its originators
regarding the operation of this process and the reactions leading to
such remarkable results. They stated:
“We have made the observation that when powdered lead sulphide (PbS),
mixed with the powdered oxide of an alkaline earth metal, _e.g._,
calcium oxide, is exposed to the action of air at bright red heat
(about 700 deg. C.), and is then allowed to cool without interrupting
the supply of air, an oxidizing decomposition takes place when dark-red
heat (about 500 deg. C.) is reached, sulphurous acid being expelled,
and a considerable amount of heat evolved; if sufficient air is then
continuously passed through the charge, dense vapors of sulphurous acid
escape, and the mixture gradually sinters together to a mass, in which
the lead of the ore is present in the form of lead oxide, provided the
air blast is continued long enough; there is no need to supply heat in
this process—the heat liberated in the reaction is quite sufficient to
keep it up.”
The inventors explained the process as follows:
“At a bright-red heat the calcium oxide (CaO) takes up oxygen from
the air supplied, forming calcium peroxide (CaO₂), which latter
afterward, in consequence of cooling down to dark-red heat, again
decomposes into monoxide and oxygen; this nascent oxygen oxidizes a
part of the lead sulphide to lead sulphate, which then reacts with a
further quantity of lead sulphide, with evolution of sulphur dioxide
and formation of lead oxide.”
Assuming the formation of calcium peroxide (CaO₂), the process
leading to the desulphurization would therefore be represented as
follows:
1. at 700° C. CaO + O = CaO₂
2. at 500° C. 4CaO₂ + PbS = 4CaO + PbSO₄
3. at the melting point PbS + PbSO₄ = 2PbO + 2SO₂ (?)
Reactions 1 and 2 combined, assuming the presence of sufficient oxygen,
give:
PbS + 4CaO + 4O = PbSO₄ + 4CaO.
Now the invention consists in applying the observation described above
to the working up of galena, and other ores containing lead sulphide,
for metallic lead; and the essential novelty of the process therefore
consists in passing air through the mass cooled to a dark-red heat (500
deg. C.).
This feature sharply distinguishes it from other known processes.
It is true that in previous processes (compare the Tarnowitz
reverberatory-furnace process, the roasting process used at
Munsterbusch near Stolberg, and others) the lead ore was mixed with
limestone or dolomite (which are converted into oxides in the early
stage of the roast) and the heat was alternately raised and lowered;
but in all cases only a surface action of the air was produced, the air
supply being provided simply by the furnace draft. Passing air through
the mass cooled down, as indicated above, leads to the important
economic advantages of reducing the fuel consumption, the losses of
lead, the manual labor (raking) and the dimensions of the roasting
apparatus.
In order to carry out the process of this invention, the powdered ore
is intimately mixed with a quantity of alkaline earth oxide, _e.g._,
calcium oxide, corresponding to its sulphur content; if the ore
already contains alkaline earth, the quantity to be added is reduced
in accordance. The mixture is heated to bright-red heat (700 deg. C.)
in the reverberatory furnace, in a strongly oxidizing atmosphere, is
then allowed to cool down to dark-red heat (500 deg. C.), also in
strongly oxidizing atmosphere, is transferred to a vessel called the
“converter,” and atmospheric air is passed through at a slight pressure
(the inventors have found a blast corresponding to 35 to 40 cm. head
of water suitable).[19] The heat liberated is quite sufficient to keep
the charge at the reaction temperature, but, if desired, hot blast may
also be used. The mixture sinters together, and (while sulphurous acid
gas escapes) it is gradually converted into a mass consisting of lead
oxide, gangue and calcium sulphate, from which the lead is extracted in
the metallic form, by any of the known methods, in the shaft furnace.
The operation is concluded as soon as the mass, by continued sintering,
has become impermeable to the blast. If the operation is properly
conducted, the gas escaping contains only small quantities of volatile
lead compounds, but on the other hand up to 8 per cent. by volume of
sulphur dioxide. This latter can be collected and further worked up.
“In place of the oxide of an alkaline earth, ferrous oxide (FeO) or
manganous oxide (MnO) may also be used.”
According to the reports on the practice of this process which have
been published,[20] conical converters of about 1700 mm. (5 ft. 6
in.) upper diameter and 1500 mm. (5 ft.) depth are used in Australian
works. At a new plant at Port Pirie (Broken Hill Proprietary Company)
converters 2400 mm. (7 ft. 10 in.) in diameter and 1800 mm. (5 ft.
11 in.) deep have been installed. These latter will hold a charge of
about eight tons. In the lower part of these converters, at a distance
of about 600 mm. (2 ft.) from the bottom, there is placed an annular
perforated plate, and upon this a short perforated tube, closed above
by a plate having only a limited number of holes.
No details have been published with regard to the European
installations. The general information which the Metallurgische
Gesellschaft[21] placed at my disposal upon request some years ago,
for use in my lecture courses, was restricted to data regarding the
consumption of fuel and labor in roasting and smelting the ores, which
was figured at about one-third or one-half of the consumption in the
former processes, to the demonstration of the large output of the
comparatively small converters, and to the reduced size of the roasting
plant as the result. But the European establishments which introduced
this process were bound by the owners of the patents, notwithstanding
the protection afforded by the patents, to give no information whatever
regarding the process to outsiders, and not to allow any inspection of
the works.
On the other hand, a great deal appeared in technical literature
which was calculated to excite curiosity. Moreover, as professor of
metallurgy, it was my duty to instruct my pupils concerning this
process among others, and it was therefore very gratifying to me
that one of the students in my laboratory took a special interest in
the treatment of lead ore. I gave him opportunity to install a small
converter, in order to carry out the process on a small scale, and
in spite of the slender dimensions of the apparatus the very first
experiments gave a complete success.
However, I could not harmonize the explanation of the process given by
the inventors with the knowledge which I had acquired in my many years’
practical experience in the manufacture of peroxides. It is clear from
the patent specification that in the roasting operation at 700 deg.
C. a compound must be formed which functions as an excellent oxygen
carrier, for on cooling to 500 deg. C. the further oxidation then
proceeds to the end not only without any external application of heat,
but even with vigorous evolution of heat. No more striking instance
than this could be desired by the theorists who have of recent years
again become so enthusiastic over the idea of catalysis. Huntington
and Heberlein regarded calcium peroxide as the oxygen carrier, but that
is a compound which cannot exist at all under the conditions which
obtain in their process. The peroxides of the alkaline earths are so
very sensitive that in preparing them the small quantities of carbon
dioxide and water must be extracted carefully from the air, and yet
in the process, in an atmosphere pregnant with carbon dioxide, water,
sulphurous acid, etc., calcium peroxide, the most sensitive of the
whole group, is supposed to form! This could not be.
The only compounds known as oxygen carriers, and capable of existing
under the conditions of the process, are calcium plumbate and plumbite.
I have emphasized this point from the first in my lectures on
metallurgy, when dealing with the Huntington-Heberlein process, and, in
point of fact, this assumption has since been proved to be correct by
the work of L. Huppertz, one of my students.
During my practical activity (1879-1891) I had prepared barium peroxide
and lead peroxide in large quantities on a manufacturing scale, the
last-mentioned through the intermediate formation of plumbites and
plumbates:
2NaOH + PbO + O = Na₂PbO₃ + H₂O
or:
4NaOH + PbO + O = Na₄PbO₄ + 2H₂O.
An experiment made in this connection showed that calcium plumbate is
formed just as readily from slaked lime and litharge as the sodium
plumbates above. Litharge is an intermediate product, produced in
large quantities in lead works, and must in any case be brought
back into the process. If, then, the litharge is roasted at a low
temperature with slaked lime, the roasting of the galena could perhaps
be entirely avoided by introducing that ore together with calcium
plumbate into the converter, after the latter had once been heated up.
Mr. Huppertz undertook the further development of this process, but I
have no information on the later experimental results, as he placed
himself in communication with neighboring lead works for the purpose
of continuing his investigation, and has not since then given me any
precise data. I will therefore confine myself to the statement that
the fundamental idea for the experiments, which Mr. Huppertz undertook
at my suggestion, was the following:
To dispense with the roasting of the galena, which is necessary
according to Huntington and Heberlein; in other words, to convert
the galena by direct blast, with the addition of calcium plumbate,
the latter being produced from the litharge which is an unavoidable
intermediate product in the metallurgy of lead and silver. (Borchers,
“Elektrometallurgie,” 3d edition, 1902-1903, p. 467.)
This alone would, of course, have meant a considerable simplification
of the roast, but the problem of the roasting of galena has been solved
in a better way by A. Savelsberg, of Ramsbeck, Westphalia, who has
determined the conditions for directly converting the galena with the
addition of limestone and water and without previous roasting. He has
communicated the following information regarding these conditions:
In order that, in blowing the air through the mixture of ore and
limestone, an alteration of the mixture may not take place owing to the
lighter particles of the limestone being carried away, it is necessary
(quite at variance with the processes in use hitherto, in which for the
sake of economy stress is laid on the precaution of charging the ore
as dry as possible into the apparatus) to add a considerable quantity
of water to the charge before introducing it into the converter. The
water serves this purpose perfectly, also preventing any change in
the mixture of ore and limestone, which invariably occurs if the ore
is used dry. The water, moreover, exerts a very beneficial action
in the process, inasmuch as it aids materially in the formation and
temporary retention of sulphuric acid, which latter then, by its
oxidizing action, greatly enhances the reaction and consequently the
desulphurization of the ore. Furthermore, the water tends to moderate
the temperature in the charge by absorbing heat in its volatilization.
In carrying out the process the converter must not be filled entirely
all at once, but first only in part, additional layers being charged
in gradually in the course of the operation. In this way a uniform
progress of the reaction in the mass is secured.
The following mode of procedure is advantageously adopted: A small
quantity of glowing fuel (coal, coke, etc.) is introduced into the
converter, which is provided at the bottom with a grate (perforated
sheet iron), the grate being first covered with a thin layer of crushed
limestone in order to protect it from the action of the red-hot coals
and ore. Upon this red-hot fuel a uniform layer of the wetted mixture
of crude ore and limestone is placed. When the surface of the first
layer has acquired a uniform red heat, a fresh layer is charged on,
and this is continued, layer by layer, until the converter is quite
full. While the layers are still being put on, the blast is passed in
at quite a low pressure, and only when the converter is entirely filled
is the whole force of the blast, at a rather greater pressure, turned
on. There then sets in a kind of slag formation, which, however, is
preceded by a very vigorous desulphurization. After the termination of
the process, which can be recognized by the fact that vapors cease to
be evolved, and that the surface of the ore becomes hard, the converter
is tipped over, and the desulphurized mass drops out as a solid cone of
slag, which is then suitably broken up for the subsequent smelting in
the shaft furnace.
Savelsberg explains the reaction of this process as follows:
“1. The particles of limestone act mechanically, gliding in between the
particles of lead ore and separating them from one another. In this
way a premature sintering is prevented, and the whole mass is rendered
loose and porous.
“2. The limestone moderates the reaction temperature produced in the
combustion of the sulphur, so that the fusion of the galena, the
formation of dust and the separation of metallic lead are avoided,
or at least kept within the limits permissible. The lowering of the
temperature of reaction is due partly to the decomposition of the
limestone into caustic lime and carbon dioxide, in which heat is
absorbed, and partly to the consumption of the quantity of heat which
is necessary in the further progress of the operation for the formation
of a slag from the gangue of the ore and the lead oxide produced.
“3. The limestone gives rise to chemical reactions. By its
decomposition it produces lime, which, at the moment of its formation,
is converted into calcium sulphate at the expense of the sulphur
in the ore. The calcium sulphate at the time of slag formation is
converted into silicate by the silica present, sulphuric acid being
evolved. The limestone therefore assists directly and forcibly in the
desulphurization of the ore, causing the formation of sulphuric acid at
the expense of the sulphur in the ore, the sulphuric acid then acting
as a strong oxidizing agent toward the sulphur in the ore.”
The most conclusive proof for the correctness of the opinion which I
expressed above, that it is very important to create at the beginning
of the operation the conditions for the formation of as much sulphate
as possible, has been furnished by Carmichael and Bradford. They
recommend that gypsum be added to the charge in place of limestone. At
one of the works of the Broken Hill Proprietary Company (where their
process has been carried on successfully, and where lead ores very rich
in zinc had to be worked up) the dehydrated gypsum was mixed with an
equal quantity of concentrate and three times the quantity of slime
from the lead ore-dressing plant, as in the table given herewith:
─────────────────┬────────┬─────────────┬──────────┬────────
│ OF THE │ OF THE │ OF THE │ OF THE
CONTENTS │ SLIME │ CONCENTRATE │ CALCIUM │ WHOLE
│ │ │ SULPHATE │ CHARGE
─────────────────┼────────┼─────────────┼──────────┼────────
Galena │ 24 │ 70 │ │ 29
Zinc blende │ 30 │ 15 │ │ 21
Pyrites │ 3 │ │ │ 2
Ferric oxide │ 4 │ │ │ 2.5
Ferrous oxide │ 1 │ │ │ 1
Manganous oxide │ 6.5 │ │ │ 5
Alumina │ 5.5 │ │ │ 3
Lime │ 3.5 │ │ 4.1 │ 10
Silica │ 23 │ │ │ 14
Sulphur trioxide │ │ │ 59 │ 12
─────────────────┴────────┴─────────────┴──────────┴────────
The charge is mixed, with addition of water, in a suitable pug-mill.
The mass is then, while still wet, broken up into pieces 50 mm. (2 in.)
in diameter, which are then allowed to dry on a floor in contact with
air; in doing so they set hard, owing to the rehydration of the gypsum.
As in the case of the Savelsberg process, the converters are heated
with a small quantity of coal, are filled with the material prepared
in the manner above described, and the charge is blown, regulating
the blast in such manner that, after the moisture present has been
dissipated, a gas of about 10 per cent. SO₂ content is produced,
which is worked up for sulphuric acid in a system of lead chambers.
The reactions are in this case the same as in the Savelsberg process,
for here also calcium sulphate is formed transitorily, which, like
other sulphates, reacts partly with sulphides, partly with silica.
Where gypsum is available and cheap, the Carmichael-Bradford process
must be given preference; in all other cases unquestionably the
Savelsberg process is superior, owing to its great simplicity.
LIME-ROASTING OF GALENA
BY W. MAYNARD HUTCHINGS
(_October 21, 1905_)
Much interest attaches to the paper by Professor Borchers, recently
presented in the _Engineering and Mining Journal_ (Sept. 2, 1905) on
“New Methods of Desulphurizing Galena,” together with an editorial on
“Lime-Roasting of Galena”; it is a curious coincidence that the same
issue contained also an article on the “Newer Treatment of Broken Hill
Sulphides,” in which is shown the importance of the new methods as a
contribution to actual practice.
For some years it had been a source of surprise to me that a new
process, so interesting and so successful as the Huntington-Heberlein
treatment of sulphide ores, should have received scarcely any notice
or discussion. This lack, however, now appears to be remedied. The
suggestion that the subject should be discussed in the _Journal_
is good, as is also that of the designation “Lime-Roasting” for a
type-name. Such observations and experiments on the subject as I have
had occasion to record have, for many years, figured in my note-books
under that heading.
Whatever may be the final results of the later processes, now before
the metallurgical world or still to come, there can be no doubt
whatever that full and exclusive credit must be given to Huntington and
Heberlein, not only for first drawing attention to the use of lime, but
also for working out and introducing practically the process. It has
been a success from the first; and so far as part of it is concerned,
it seems to be an absolute and fundamental necessity which later
inventors can neither better nor set aside. The other processes, since
patented, however good they may be, are simply grafts on this parent
stem.
It is, however, quite certain that Huntington and Heberlein, in the
theoretical explanation of the process, failed to understand the most
important reactions. Their attributing the effect to the formation and
action of calcium peroxide affords a sad case of _a priori_ assumption
devoid of any shred of evidence. As Professor Borchers points out,
calcium peroxide, so difficult to produce and so unstable when formed,
is an absolute and absurd impossibility under the conditions in
question. Probably many rubbed their eyes with astonishment on reading
that part of the patent on its first appearance, and hastened to look
up the chemical authorities to refresh their minds, lest something as
to the nature of calcium peroxide might have escaped them.
Fortunately the patent law is such that there was no danger of a really
good and sound invention being invalidated by a wrong theoretical
explanation by its originators. But, nevertheless, it was a misfortune
that the inventors did not understand their own process. Had they
known, they could have added a few more words to their patent-claims
and rendered the Carmichael patent an impossibility.
Professor Borchers appears to consider that the active agent in the new
process is calcium plumbate. That this compound may play a part at some
stage of the process may be true; this long ago suggested itself to
some others. We may yet expect to hear that the experiments undertaken
by Professor Borchers himself, and by others at his instigation (in
which calcium plumbate is separately prepared and then brought into
action with lead sulphide), have given good results. But it does not
appear so far that there is any real proof that calcium plumbate is
formed in the Huntington-Heberlein or other similar processes; and it
is difficult to see at what stage or how it would be produced under the
conditions in question. This is a point which research may clear up,
but it should not be taken for granted at this stage. Indeed, it seems
to me that the results obtained may be fairly well explained without
calling calcium plumbate into play at all.
Of course the action of lime in contact with lead sulphide excited
interest many years before the new process came into existence. My own
attention to it dates back more than a dozen years before that time (I
was in charge of works where I found the old “Flintshire process” still
in use).
Percy pointed out, in his work on lead smelting, that on the addition
of slaked lime to the charge, at certain stages, to “stiffen it up,”
the mixture could be seen to “glow” for a time. When I myself saw this
phenomenon, I commenced to make some observations and experiments.
Also (as others probably had done), I had observed that charges of lead
with calcareous gangue are roasted more rapidly and better than others,
and to an extent which could not be wholly explained by simple physical
action of the lime present.
Simple experiments made in assay-scorifiers in a muffle, on lime
roasting, are very striking, and I think quite explain a good part
of what takes place up to a certain stage in the processes now under
consideration. I tried them a number of years ago, on many sorts of
ore, and again more recently, when studying the working of the new
patents. For illustration, I will take one class of ore (Broken Hill
concentrate), using a sample assaying; Pb, 58 per cent.; Fe, 3.6 per
cent.; S, 14.6 per cent.; SiO₂, 3 per cent. The ore contained some
pyrite. If two scorifiers are charged, one with the finely powdered
ore alone, and one with the ore intimately mixed with, say, 10 per
cent. of pure lime, and placed side by side just within a muffle at
low redness, the limed charge will soon be seen to “glow.” Before the
simple ore charge shows any sign of action, the limed charge rapidly
ignites all over, like so much tinder, and heats up considerably above
the surrounding temperature, at the same time increasing noticeably
in bulk. This lasts for some time, during which hardly any SO₂
passes off. After the violent glowing is over, the charge continues
to calcine quietly, giving off SO₂, but is still far more active
than its neighbor. If, finally, the fully roasted charge is taken out,
cooled and rubbed down, it proves to contain no free lime at all, but
large quantities of calcium sulphate can be dissolved out by boiling in
distilled water. For instance, in one example where weighed quantities
were taken of lime and the ore mentioned, the final roasted material
was shown to contain nearly 23 per cent. of CaSO₄; the quantity
actually extracted by water was 20.2 per cent. Further tests show
that the insoluble portion still contains calcium sulphate intimately
combined with lead sulphate, but not extractable by water.
There is no doubt that when lead sulphide (or other sulphide) is
heated with lime, with free access of air, the lime is rapidly and
completely converted into sulphate. The strong base, lime, apparently
plays the part of “catalyzer” in the most vigorous manner, the first
SO₂ evolved being instantly oxidized and combined with the lime
to sulphate, with so strong an evolution of heat that the operation
spreads rapidly and still goes on energetically, even if the scorifier
is taken out of the muffle. Also, the “catalytic” action starts the
oxidation of the sulphides at a far lower temperature than is required
when they are roasted alone.
If, in place of lime, we take an equivalent weight of pure calcium
carbonate and intimately mix it with ore, we obtain just the same
action, only it takes a little longer to start it. Once started, it
is almost as vigorous and rapid, and with the same results. It does
not seem correct to assume (as is usually done) that the carbonate has
first to be decomposed by heat, the lime then coming into action. The
reaction commences in so short a time and while the charge is still
so cool, that no appreciable driving off of CO₂ by heat only can
have taken place. The main liberation of the CO₂ occurs during the
vigorous exothermic oxidation of the mixture, and is coincident with
the conversion of the CaO into CaSO₄.
If, in place of lime or its carbonate, we use a corresponding quantity
of pure calcium sulphate and mix it with the ore, we see very energetic
roasting in this case also, with copious evolution of sulphur dioxide,
only it is much more energetic and rapid and occurs at a lower
temperature than in the case of a companion charge of ore alone.
It is very easily demonstrated that the CaSO₄ in contact with the
still unoxidized ore (whether it has been introduced ready made or has
been formed from lime or limestone added) greatly assists the further
roasting, in acting as a “carrier” and enabling calcination to take
place more rapidly and easily and at a lower temperature than would
otherwise be the case.
The result of these experiments (whether we mix the ore with CaO,
CaCO₃, or CaSO₄) is that we arrive with great ease and rapidity
at a nearly dead-sweet roast. The lime is converted into sulphate, and
the lead partly to sulphate and partly to oxide. Two examples out of
several, both from the above ore, gave results as follows:
No. 1—Roasted with 20 per cent. CaCO₃ (= 11.2 per cent. CaO);
sulphide sulphur, 0.02 per cent.; sulphate sulphur, 9.30 per cent.;
total sulphur, 9.32 per cent.
No. 8—Roasted with 27.2 per cent. CaSO₄ (= 11 per cent. CaO);
sulphide sulphur, 0.05 per cent.; sulphate sulphur, 11.28 per cent.;
total sulphur, 11.33 per cent.
If these calcined products are now intimately mixed with additional
silica (in about the proportions used in the Huntington-Heberlein
process) and strongly heated, fritting is brought about and the sulphur
content is reduced by the decomposition of the sulphates by the silica.
Thus, the resultant material of experiment No. 1, above, when treated
in this manner with strong heat for three hours, was sintered to a mass
which was quite hard and stony when cold, and which contained 6.75
per cent. of total sulphur. Longer heating drives out more sulphur,
but a very long time is required; in furnaces, and on a large scale,
it is with great difficulty and cost that a product can be obtained
comparable with that which is rapidly and cheaply turned out from the
“converters” of the new process.
To return to the Huntington-Heberlein process, working, for example,
on an ore more or less like the one given above, we may assume that,
during the comparatively short preliminary roast, the lime is all
rapidly converted into CaSO₄ and that some PbSO₄ is also formed
(but not much, as the mixture to be transferred from the furnace to
the converter requires not less than 6 to 8 per cent. of sulphur to
be still present as sulphide, in order that the following operation
may work at its best). As the blast permeates the mass, oxidation is
energetic; no doubt that CaSO₄ here plays a very important part
as a carrier of oxygen, in the same manner as we can see it act on a
scorifier or on the hearth of a furnace.
What the later reactions are does not seem so clear. They are quite
different from those on the scorifier or on the open hearth of a
furnace, and result in the rapid formation (in successive layers of
the mixture, from the bottom upward) of large amounts of lead oxide,
fluxing the silica and other constituents to a more or less slaggy
mass, which decomposes the sulphates and takes up the CaO into a
complex and easily fused silicate. It is true that, as a whole, the
contents of a well-worked converter are never very hot, but locally
(in the regions where the progressive reaction and decomposition from
below upward is going on) the temperature reached is considerable. This
formation of lead oxide is so pronounced at times that one may see in
the final product considerable quantities of pure uncombined litharge.
When the work is successful, the mass discharged from the converters
is a basic silicate of PbO, CaO, and oxides of other metals present,
and nearly all the sulphates have disappeared. A large piece of yellow
product (which was taken from a well-worked converter) contained only
1.1 per cent. of total sulphur.
It may be that calcium plumbate is formed and plays a part in these
reactions; but its presence would be difficult to prove, and its
formation and existence during these stages would not be easy to
explain. Neither does it seem necessary, as the whole thing appears to
be capable of explanation without it.
While the mixture in the converter is still dry and loose, energetic
oxidation of the sulphides goes on, with the intervention of the
CaSO₄ as a carrier. As soon as the heat rises sufficiently, fluxing
commences in a given layer and sulphates are decomposed. The liberated
sulphuric anhydride, at the locally high temperature and under the
existing conditions, will act with the greatest possible vigor on the
sulphides in the adjacent layers; these layers will then in their
turn flux and act on those above them, till the whole charge is
worked out. The column of ore is of considerable hight, requiring a
blast of 1½ lb., or perhaps more, in the larger converters now used.
This pressure of the oxidizing blast (and of the far more powerfully
oxidizing sulphuric anhydride, continuously being liberated within the
mass of ore, locally very hot) constitutes a totally different set of
conditions from those obtained on the hearth of a furnace with the ore
in thin layers, where it is neither so hot nor under any pressure.
It is to these conditions, in which we have the continued intense
action of red-hot sulphuric anhydride under a considerable pressure
(together with the earlier action of the CaSO₄), that the remarkable
efficiency of the process seems to me to be due.
In the Carmichael process, the preliminary roast is done away with,
CaSO₄ being added directly instead of having to be formed during the
operation from CaO and the oxidized sulphur of the ore. The charge in
the converter has to be started by heat supplied to it, and the work
then goes forward on the same lines as in the Huntington-Heberlein
process, so that we may assume that the reactions are the same and come
under the same explanation.
Carmichael was quick to see what was really an important part and a
correct explanation of the original process. He was not misled by wrong
theory about any mythical calcium peroxide, and so he obtained his
patent for the use of CaSO₄ and the dispensing of the roast in a
furnace.
This process would always be limited in its application by the
comparative rarity of cheap supplies of gypsum, but it appears to be
a great success at Broken Hill; there it is not only of importance in
working the leady ores, but also for making sulphuric acid for the new
treatment of mixed sulphides by the Delprat and Potter methods. For
this purpose, the use of CaSO₄ will have the additional advantage
that the mixture to be worked in the converter will contain not
only the sulphur of the ore, but also that of the added gypsum; on
decomposition, it will yield stronger gases for the lead chambers of
the acid plant.
Finally comes the Savelsberg patent, which is the simplest of all;
not only (like the Carmichael process) avoiding the preliminary roast
with its extra plant, but also not requiring the use of ready-made
CaSO₄, as it uses raw ore and limestone directly in the converter.
I have no knowledge as to actual results of this process; and, so
far as I am aware, nothing on the subject has been published. But
Professor Borchers evidently has some information about it, and
regards it as the most successful of the methods of carrying out the
new ideas. On the face of it, there seems no reason why it should not
attain all the results desired, as the chemical and physical actions
of the CaO, and of the CaSO₄ formed from it, should come into play
in the same manner and in the same order as in the original process;
as it is carried out in the identical converter used by Huntington
and Heberlein, the final reactions (as suggested above) will take
place under the same conditions as to continuous decomposition _under
considerable heat and pressure_, which I regard as the most vital part
of the whole matter.
It is well to emphasize again the fact that the idea, and the means of
obtaining these vital conditions, owe their origination to Huntington
and Heberlein.
THEORETICAL ASPECTS OF LEAD-ORE ROASTING[22]
BY C. GUILLEMAIN
(March 10, 1906)
It is well known that the process of roasting lead ores in
reverberatory furnaces proceeds in various ways according to the
composition of the ore in question. Thus in roasting a sulphide lead
ore rich in silica, one of the reactions is:
PbS + 3O = PbO + SO₂.
But this reaction is incomplete, for the gases which pass on in the
furnace are rich in SO₂ and in SO₃. And so it is found that
whatever lead oxide is formed passes over almost immediately into lead
sulphate, according to the reaction:
PbO + SO₂ + O = PbSO₄.
This reaction is the chief one which takes place. Whether the silicious
gangue serves as a catalyzer for the sulphur dioxide, or whether it
serves merely to keep the galena open to the action of the gases, the
end result of the roast is usually the formation of lead sulphate
according to the above reaction.
In the case of an ore rich in galena, a slow roast is essential, for it
is desired to have the following reaction take place during the latter
part of the roast:
PbS + 3PbSO₄ = 4PbO + 4SO₂.
Now, if the heating were too rapid, not enough lead sulphate would be
found to react with the unaltered galena. The quick roasting of a rich
ore would result in the early sintering of the charge, and sintering
prevents the further formation of lead sulphate. Whether this sintering
(which takes place so easily and which is so harmful in the latter part
of the process) is due to the low melting point of the lead sulphide,
whether the heat evolved by the reaction
PbS + 3O = PbO + SO₂
is sufficient to melt the lead sulphide, or whether other
thermochemical effects (notably the preliminary sulphatizing of the
lead sulphide) come into play, must for the present be undecided.
Suffice it to say that the sintering of the charge works against a good
roast.
In the Tarnowitz process a definite amount of lead sulphide is
converted into lead sulphate by a preliminary roast. The sulphate then
reacts with the unaltered lead sulphide, and metallic lead is set free,
thus:
PbS + PbSO₄ = 2Pb + 2SO₂.
But when a very little of the sulphide has been transformed into
sulphate, and when there is so little of the latter present that only
a small amount of lead sulphide can be reduced to metallic lead, the
mass of ore begins to sinter and grow pasty. Very little lead could be
formed were it not for the addition of crushed lime to the charge just
before the sintering begins. This lime breaks up the charge and cools
it, prevents any sintering, and allows the continued formation of lead
sulphate.
It scarcely can be held that the lime has any chemical effect in
forming lead sulphate, or in forming a hypothetical compound of lead
and calcium. Even if such theories were tenable from a physico-chemical
point of view, they would be lessened in importance by the fact that
other substances, such as purple ore or puddle cinder, act just as well
as the lime.
There are now to be mentioned several new processes of lead-ore
roasting whose operations fall so far outside the common ideas on
the subject that their investigation is full of interest. For a long
time the attempt had been made to produce lead directly by blowing
air through lead sulphide in a manner analogous to the production of
bessemer steel or the converting of copper matte. In the case of the
lead sulphide, the oxidation of the sulphur was to furnish the heat
necessary to carry on the process.
After many attempts along this line, Antonin Germot has perfected a
method wherein, by blowing air through molten galena, metallic lead
is obtained.[23] About 60 per cent. of a previously melted charge of
galena is sublimed as lead sulphide, and the rest remains behind as
metallic lead. The disadvantages of the process are the difficulties of
collecting all of the sublimate and of working it up. Moreover, it is
impossible as yet to secure two products of which one is silver-free
and the other silver-bearing. The silver values are in both the
metallic lead and in the sublimed lead sulphide.
While the process just described answers for pure galena, it fails
with ores which contain about 10 per cent. of gangue. In the case of
such ores, they form a non-homogeneous mass when melted, and the blast
penetrates the charge with difficulty. If the pressure is increased
the air forces itself out through tubes and canals which it makes for
itself, and the charge freezes around these passages.
Messrs. Huntington and Heberlein have gone a little farther. Although
they are unable to obtain metallic lead directly, they prepare the ore
satisfactorily for smelting in the blast furnace, after their roasting
is completed. The inventors found that if lead sulphide is mixed with
crushed lime, heated with access of air, and then charged into a
converter and blown, the sulphur is completely removed in the form of
sulphur dioxide. The charge, being divided by the lime, remains open
uniformly to the passage of air, and sinters only when the sulphur is
eliminated.
The inventors announce, as the theory of their process, that at 700
deg. C. the lime forms a dioxide of calcium (CaO₂) which at 500 deg.
C. breaks down into lime (CaO) and nascent oxygen. This nascent oxygen
oxidizes the lead sulphide to lead sulphate according to the reaction:
PbS + 4O = PbSO₄.
Furthermore it is claimed that the heat evolved by this last reaction
is large enough to start and keep in operation a second reaction, namely
PbS + PbSO₄ = 2PbO + 2SO₂.
The theory, as just mentioned, cannot be accepted, and some of the
reasons leading to its rejection will be given.
It is well established that the simple heating of lime with access of
air will not result in further oxidation of the calcium. The dioxide
of calcium cannot be formed even by heating lime to incandescence in
an atmosphere of oxygen, nor by fusing lime with potassium chlorate.
Moreover, calcium stands very near barium in the periodic system. And
as the dioxide of barium is formed at a low temperature and breaks
up on continued heating, it seems absurd to suppose that the dioxide
of calcium would act in exactly the opposite manner. Moreover, a
consideration of the thermo-chemical effects will disclose more
inconsistencies in the ideas of the inventors. The breaking up of
CaO₂ into CaO and O is accompanied by the evolution of 12 cal. The
reaction of the oxygen (thus supposed to be liberated) upon the lead
sulphide is strongly exothermic, giving up 195.4 cal. So much heat is
produced by these two reactions that, if the ideas of the inventors
were true, the further breaking up of the calcium dioxide would stop,
as the whole charge would be above 500 deg. C. It appears, then, that
the explanations suggested by Messrs. Huntington and Heberlein are
untrue.
In the usual roasting process, as carried out in reverberatory
furnaces, it is well established that the gangue, and whatever
other substances are added to the ore, prevent mechanical locking
up of charge particles, since they stop sintering. It is not at all
improbable that in the new roasting process the chief, if not the only,
part played by the lime is the same as that played by the gangue in
reverberatory-furnace roasting. A few observations leading to this
belief will be given.
It is known that other substances will answer just as well as lime
in this new roasting process. Such substances are manganese and iron
oxides. Not only these two substances, but in fact any substance which
answers the purpose of diminishing the local strong evolution of heat,
due to the reaction:
PbS + 3O = PbO + SO₂,
serves just as well as the lime. This fact is proved by exhaustive
experiments in which mixtures of lead sulphide on the one hand, and
quartz, crushed lead slags, iron slags, crushed iron ores, crushed
copper slags, etc., on the other hand, were used for blowing. All
these substances are such that any chemical action, analogous to the
splitting up of CaO₂, or the formation of plumbates as suggested
by Dr. Borchers, cannot be imagined. The time is not yet ripe, without
more experiments on the subject, to assert conclusively that there
is no acceleration of the process due to the formation of plumbates
through the agency of lime. But the facts thus far secured point out
that such reactions are, at least, not of much importance.
Theoretical considerations point out that it ought to be possible to
avoid the injurious local increase of temperature during the progress
of this new roasting process, without having to add any substance
whatever. To explain: The first reaction taking place in the roasting is
PbS + 3O = PbO + SO₂ + 99.8 cal.
Now the heat thus liberated may be successfully dispersed if there is,
in simultaneous progress, the endothermic reaction:
PbS + 3PbSO₄ = 4PbO + 4SO₂ - 187 cal.
Hence if there could be obtained a mixture of lead sulphide and of
lead sulphate in the proportions demanded by the above reaction, then
such a mixture ought to be blown successfully to lead oxide without
the addition of any other substance. Such a process has, in fact, been
carried out. The original galena is heated until the required amount of
lead sulphate has been formed. Then the mixture of lead sulphide and
of lead sulphate is transferred to a converter and blown successfully
without the addition of any other substance.
The adaptability of an ore to the process just mentioned depends on
the cost of the preliminary roast and the thoroughness with which it
must be done. As is known, when lead sulphide is heated with access of
air, it is very easy to form sintered incrustations of lead sulphate.
If these incrustations are not broken up, or if their formation is not
prevented by diligent rabbling, the further access of air to the mass
is prevented and the oxidation of the charge stops. If ores with such
incrustations are placed in the converter without being crushed, they
remain unaltered by the blowing. If the incrustations are too numerous
the converting becomes a failure.
It has been found that the adoption of mechanical roasting furnaces
prevents this. Such furnaces appear to stop the frequent failures of
the blowing which are due to the lack of care on the part of the
workmen during the preliminary roasting. Moreover, in such mechanical
furnaces a more intimate mixture of the sulphide with the sulphate
is obtained, and the degree of the sulphatizing roast is more easily
controlled.
As a summary of the facts connected with this new blowing process, it
may be stated that the best method of working can be determined upon
and adopted if one has in mind the fact that the amount of substance
(lime) to be added is dependent on: 1, the amount of sulphur present;
2, the forms of oxidation of this sulphur; 3, the amount of gangue
in the ore; 4, the specific heats of the gangue and of the substance
added; 5, the degree of the preparatory roasting and heating.
For example, with concentrates which run high in sulphur, there is
required either a large amount of additional material, or a long
preliminary roast. The specific heat of the added material must be
high, and the heat evolved by the oxidation of the sulphur in the
preliminary roast must be dispersed. Oftentimes it is necessary to cool
the charge partially with water before blowing. On the other hand, if
the ore runs low in sulphur, the preliminary roast must be short, and
the temperature necessary for starting the blowing reactions must be
secured by heating the charge out of contact with air. Not only must no
flux be added, but oftentimes some other sulphides must be supplied in
order that the blowing may be carried out at all.
The opportunity for the acquisition of more knowledge on this subject
is very great. It lies in the direction of seeing whether or not the
strong local evolution of heat cannot be reduced by blowing with gases
poor in oxygen rather than with air. Mixtures of filtered flue gases
and of air can be made in almost any proportion, and such mixtures
would have a marked effect upon the possibility of regulating the
progress of the oxidation of the various ores and ore-mixtures which
are met with in practice.
METALLURGICAL BEHAVIOR OF LEAD SULPHIDE AND CALCIUM SULPHATE[24]
BY F. O. DOELTZ
(January 27, 1906)
In his British patent,[25] for desulphurizing sulphide ores, A. D.
Carmichael states that a mixture of lead sulphide and calcium sulphate
reacts “at dull red heat, say about 400 deg. C.,” forming lead sulphate
and calcium sulphide, according to the equation:
PbS + CaSO₄ = PbSO₄ + CaS.
Judging from thermo-chemical data, this reaction does not seem
probable. According to Roberts-Austen,[26] the heats of formation (in
kilogram-calories) of the different compounds in this equation are as
follows: PbS = 17.8; CaSO₄ = 318.4; PbSO₄ = 216.2; CaS = 92.
Hence we have the algebraic sum:
-17.8 - 318.4 + 216.2 + 92 = -28.0 cal.
As the law of maximum work does not hold, experiment only can
decide whether this decomposition takes place or not. The following
experiments were made:
_Experiment 1._—Coarsely crystalline and specially pure galena was
ground to powder. Some gypsum was powdered, and then calcined. The
powdered galena and calcined gypsum were mixed in molecular proportions
(PbS + CaSO₄), and heated for 1½ hours to 400 deg. C., in a stream
of carbon dioxide in a platinum resistance furnace. The temperature was
measured with a Le Chatelier pyrometer. The material was allowed to
cool in a current of carbon dioxide.
The mixture showed no signs of reaction. Under the magnifying glass the
bright cube-faces of galena could be clearly distinguished. If any
reaction had taken place, in accordance with the equation given above,
no bright faces of galena would have remained.
_Experiment 2._—A similar mixture was slowly heated, also in the
electric furnace, to 850 deg. C., in a stream of carbon dioxide, and
was kept at this temperature for one hour.
It was observed that some galena sublimed without decomposition, being
redeposited at the colder end of the porcelain boat (7 cm. long), in
the form of small shining crystals. The residue was a mixture of dark
particles of galena and white particles of gypsum, in which no evidence
of any reaction was visible under the microscope. That galena sublimes
markedly below its melting point has already been noted by Lodin.[27]
_Experiment 3._—In order to determine whether the inverse reaction
takes place, for which the heat of reaction is + 28.0 cal., the
following equations are given:
PbSO₄ + CaS = PbS + CaSO₄;
- 216.2 - 92 + 17.8 + 318.4 = 28.
A mixture of lead sulphate and calcium sulphide was heated in a
porcelain crucible in a benzine-bunsen flame (Barthel burner). The
materials were supplied expressly “for scientific investigation” by the
firm, C. A. F. Kahlbaum.
The white mixture turned dark and presently assumed the color which
would correspond to its conversion into lead sulphide and calcium
sulphate. This experiment is easy to perform.
_Experiment 4._—The same materials, lead sulphate and calcium sulphide,
were mixed in molecular ratio (PbSO₄ + CaS), and were heated for 30
minutes to 400 deg. C., on a porcelain boat in the electric furnace,
in a current of carbon dioxide. The mixture was allowed to cool in a
stream of carbon dioxide, and was withdrawn from the furnace the next
day (the experiment having been made in the evening).
The mixture showed a dark coloration, similar to that of the last
experiment; but a few white particles were still recognizable. The
material in the boat smelled of hydrogen sulphide.
_Experiment 5._—A mixture of pure galena and calcined gypsum, in
molecular ratio (PbS + CaSO₄), was placed on a covered scorifier
and introduced into the hot muffle of a petroleum furnace, at 700 to
800 deg. C. The temperature was then raised to 1100 deg. C.
From 5 g. of the mixture a dark-gray porous cake weighing 3.7g. was
thus obtained. There was some undecomposed gypsum present, recognizable
under the magnifying glass. No metallic lead had separated out. When
hot hydrochloric acid was poured over the mixture, it evolved hydrogen
sulphide. The fracture of the cake showed isolated shining spots. The
supposition that it was melted or sublimed galena was confirmed by
the aspect of the cake when cut with a knife; the surface showed the
typical appearance of the cut surface of melted galena. On cutting, the
cake was found to be brittle, with a tendency to crumble. On boiling
with acetic acid, a little lead went into solution. Wetting with water
did not change the color of the crushed cake.
_Experiment 6._—In his experiments for determining the melting point
of galena, Lodin[28] found that, in addition to its sublimation at a
comparatively low temperature, the galena also undergoes oxidation if
carbon dioxide is used as the “neutral” atmosphere. Lodin was therefore
compelled to use a stream of nitrogen in his determination of the
melting point of galena. Now the temperature of experiment 2 (850 deg.
C.), described heretofore, is not as high as the melting point of
galena (which lies between 930 and 940 deg. C.); therefore experiment 2
was repeated in a stream of nitrogen, so as to insure a really neutral
atmosphere. A mixture of galena and calcined gypsum in molecular
ratio (PbS + CaSO₄) was heated to 850 deg. C., was kept at this
temperature for one hour, and allowed to cool, the entire operation
being carried out in a stream of nitrogen.
Again, galena had sublimed away from the hotter end of the porcelain
boat (6.5 cm. long), and had been partially deposited in the form of
small crystals of lead sulphide at the colder end. The material in
the boat consisted of a mixture of particles having the dark color
of galena, and others with the white color of gypsum, the original
crystals of gypsum and the bright surfaces of the lead sulphide being
distinctly recognizable under the magnifying glass. The loss in weight
was 1.9 per cent.
_Experiment 7._—For the same reason as in 2, experiment 5 was also
repeated, using a current of nitrogen. A mixture of galena and
calcined gypsum, in molecular ratio (PbS + CaSO₄) was heated in a
porcelain boat to 1030 deg. C., in a platinum-resistance furnace, and
allowed to cool, being surrounded by a stream of nitrogen during the
whole period.
Some sublimation of lead sulphide again took place. The mixture was
seen to consist of white particles of gypsum, and others dark, like
galena. The loss in weight was 3.5 per cent. The mixture had sintered
together slightly; with hot hydrochloric acid, it evolved hydrogen
sulphide. On boiling with acetic acid, a little lead (only a trace)
went into solution. There was, therefore, practically no lead oxide
present; no metallic lead had separated out.
_Experiment 8._—In experiment 3, lead sulphate and calcium sulphide
were mixed roughly and by hand (i.e., not weighed out in molecular
ratio); in this experiment such a mixture of lead sulphate and calcium
sulphide in molecular ratio (PbSO₄ + CaS) was heated in a porcelain
crucible in a benzine-bunsen flame. It presently turned dark, and a
dark gray product was obtained, as in the former experiment.
_Experiment 9._—In a mixture of lead sulphate and sodium sulphide in
molecular ratio (PbSO₄ + Na₂S), the constituents react directly
on rubbing together in a porcelain mortar. The mass turns dark gray,
with formation of lead sulphide and sodium sulphate.
If a similar mixture is heated, it also turns dark gray. On lixiviation
with water, a solution is obtained which gives a dense white
precipitate with barium chloride.
_Experiment 10._—If lead sulphate and calcium sulphide are rubbed
together in a mortar, the mass turns a grayish-black.
_Conclusion._—From these experiments I infer that the reaction
PbS + CaSO₄ = PbSO₄ + CaS
does not take place, but, on the contrary, that when lead sulphate and
calcium sulphide are brought together, the tendency is to form lead
sulphide and calcium sulphate.
Nevertheless, on heating a mixture of galena and gypsum in contact with
air, lead sulphate will be formed along with lead oxide; not, however,
owing to any double decomposition of the galena with the gypsum, but
rather to the formation of lead sulphate from lead oxide and sulphuric
acid produced by catalysis, thus:
PbO + SO₂ + O = PbSO₄.
This is the well-known process which always takes place in roasting
galena, the explanation of which was familiar to Carl Friedrich
Plattner. That the presence of gypsum has any chemical influence on
this process seems to be out of the question according to the above
experiments.
THE HUNTINGTON-HEBERLEIN PROCESS
BY DONALD CLARK
(October 20, 1904)
The process was patented in 1897, and is based on the fact that galena
can be desulphurized by mixing it with lime and blowing a current of
air through the mixture. If the temperature is dull red at the start,
no additional source of heat is necessary, because the reaction causes
a great rise in temperature. The chemistry of the process cannot be
said at present to have been worked out in detail.
The reactions given by the patentees are not satisfactory, since
calcium dioxide is formed only at low temperatures and is readily
decomposed on gently warming it; lead oxide, however, combines with
oxygen under suitable conditions at a temperature not exceeding 450
deg. C. and forms a higher oxide, and it is probable that this unites
with the lime to form calcium plumbate. The reaction between sulphides
and lime when intimately mixed and heated may be put down as
CaO + PbS = CaS + PbO.
In contact with the air the calcium sulphide oxidizes to sulphite, then
to sulphate, then reacts with lead oxide, giving calcium plumbate and
sulphur dioxide,
CaSO₄ + PbO = CaPbO₃ + SO₂.
Further, calcium sulphate will also react with galena, giving calcium
sulphide and lead sulphate; the calcium sulphide is oxidized, by air
blown through, to calcium sulphate again, the ultimate reaction being
CaSO₄ + PbS + O = CaPbO₃ + SO₂.
In all cases the action is oxidizing and desulphurizing. It was found
that oxides of iron and manganese will, to a certain extent, serve the
same purpose as lime, and on application to complex ores, especially
those containing much blende, that these may be desulphurized as well
as galena. In the case of zinc sulphide the decomposition is probably
due to the interaction of sulphide and sulphate.
ZnS + 3ZnSO₄ = 4ZnO + 4SO₂.
The process has now been adopted by the Broken Hill Proprietary
Company at its works at Port Pirie, the Tasmanian Smelting Company,
Zeehan, the Fremantle Smelting Works, West Australia, and the Sulphide
Corporation’s works at Cockle Creek, New South Wales.
The operations carried on at the Tasmania Smelting Works comprise
mixing pulverized limestone, galena and slag-making materials and
introducing the mixture either into hand-rabbled reverberatories or
mechanical furnaces with rotating hearths. After a roast, during
which the materials have become well mixed and most of the limestone
converted into sulphate and about half of the sulphur expelled, the
granular product is run while still hot into the Huntington-Heberlein
converters. These consist of inverted sheet-iron cones, hung on
trunnions, the diameter being 5 ft. 6 in. and the depth 5 ft. A
perforated plate or colander is placed as a diaphragm across the apex
of the cone, the small conical space below serving as a wind-box into
which compressed air is forced. A hood above the converter serves to
carry away waste gases. As soon as the vessel is filled, air under a
pressure of 17 oz. is forced through the mass, which rapidly warms up,
giving off sulphur dioxide abundantly. The temperature rises and the
mixture fuses, and in from two to four hours the action is complete.
The sulphur is reduced from 10 to 1 per cent., and the whole mass is
fritted and fused together. The converter is emptied by inverting it,
when the sintered mass falls out and is broken up and sent to the
smelters. There are 12 converters, of the size indicated, for the two
mechanical furnaces, of 15 ft. diameter. Larger converters of the
same type were erected to deal with the product from the hand-rabbled
roasters.
At Cockle Creek, New South Wales, the galena concentrate is reduced
to 1.5 mm., more than 60 per cent. of the material being finer; the
limestone is crushed down to from 10 to 16 mesh; silica is also added,
if it does not exist in the ore, so that, excluding the lead, the rest
of the bases will be in such proportion as to form a slag running about
20 per cent. silica. The mixture may contain from 25 to 50 per cent.
lead, and from 6 to 9 per cent. lime; if too much lime is added the
final product is powdery, instead of being in a fused condition. This
is given a preliminary roast in a Godfrey furnace.
The Godfrey furnace is characterized by a rotating, circular hearth
and a low dome-shaped roof. Ore is fed through a hopper at the center
and deflected outward by blades attached to a fixed radial arm. At
each revolution the ore is turned over and moved outward, the mount of
deflection of the blades, which are adjustable, and rate of rotation of
the hearth, determining the output.
The hot semi-roasted ore is discharged through a slot at the
circumference of the roaster. This may contain from 12 to 6.5 per
cent. of sulphur, but from 6.5 to 8 per cent. is held to be the most
suitable quantity for the subsequent operations. Thorough mixing is of
the utmost importance, for if this is not done the mass will “volcano”
in the converter; that is, channels will form in the mass through which
the gases will escape, leaving lumps of untouched material alongside.
The action can be started if a little red-hot ore is run into the
converter and cold ore placed above it; the whole mass will become
heated up, and the products will fuse, and sinter into a homogeneous
mass showing none of the original ingredients. At Cockle Creek the time
taken is stated to be five hours; a small air-pressure is turned on at
first, and ultimately it is increased to 20 oz.
Operations at Port Pirie are conducted on a much larger scale. A
mixture of pulverized galena, powdery limestone, ironstone and sand
is fed into Ropp furnaces, of which there are five, by means of a
fluted roll placed at the base of a hopper. Each roaster deals with
100 tons of the mixture in 24 hours. About 50 per cent. of the sulphur
is eliminated from the ore by the Ropps (the galena in this case being
admixed with a large amount of blende, there being only 55 per cent.
of lead and 10 per cent. of zinc in the concentrate produced at the
Proprietary mine). The hot ore from the roasters is trucked to the
converters, there being 17 of these ranged in line. The converters here
are large segmental cast-iron pots hung on trunnions; each is about 8
ft. diameter and 6 ft. deep, and holds an 8-ton charge. At about two
feet from the bottom an annular perforated plate fits horizontally;
a shallow frustrum of a cone, also perforated, rests on this; while
a plate with a few perforations closes the top of the frustrum. The
whole serves as a wind-box. A conical hood with flanged edges rests
on the flanged edges of the converter, giving a close joint. This
hood is provided with doors which allow the charge to be barred if
necessary. A pipe about 1 ft. 9 in. diameter, fitted with a telescopic
sliding arrangement, allows for the raising or lowering of the hood by
block and tackle, and thus enables the converter to be tilted up and
its products emptied. The cast-iron pots stand very well; they crack
sometimes, but they can be patched up with an iron strap and rivets.
Only two pots have been lost in 18 months.
Air enters at a pressure of about 24 oz. and the time taken for
conversion is about four hours. The sulphur contents are reduced to
about three per cent. It is found that the top of the charge is not so
well converted as the interior. There is practically no loss of lead
or silver due to volatilization and very little due to escape of zinc.
It has also been found that practically all the limestone fed into
the Ropp is converted into calcium sulphate; also that a considerable
portion of lead becomes sulphate, and it is considered that lead
sulphate is as necessary for the process as galena.
The value of the process may be judged from the fact that better work
is now done with 8 blast furnaces than was done with 13 before the
process was adopted. In addition to the sintered product from the
Huntington-Heberlein pots, sintered slime, obtained by heap roasting,
and flux consisting of limestone and ironstone, are fed into the
furnaces, which take 2000 long tons per day of ore, fluxes and fuel.
The slags now being produced average: SiO₂, 25 to 26 per cent.; FeO,
1 to 3 per cent.; MnO, 5 to 5.5; CaO, 15.5 to 17; ZnO, 13; Al₂O₃,
6.5; S, 3 to 4; Pb, by wet assay, 1.2 to 1.5 per cent.; and Ag, 0.7 oz.
per ton. Although this comparatively large quantity of sulphur remains,
yet no matte is formed.
THE HUNTINGTON-HEBERLEIN PROCESS AT FRIEDRICHSHÜTTE[29]
BY A. BIERNBAUM
(September 2, 1905)
Nothing, for some time past, has caused such a stir in the
metallurgical treatment of lead ores, and produced such radical
changes at many lead smelting works, as the introduction of the
Huntington-Heberlein process. This process (which it may be remarked,
incidentally, has given rise to the invention of several similar
processes) represents an important advance in lead smelting, and,
now that it has been in use for some time at the Friedrichshütte,
near Tarnowitz, in Upper Silesia, and has there undergone further
improvement in several respects, a comparison of this process with the
earlier roasting process is of interest.
At the above-mentioned works, up to 1900 the lead ore was
treated exclusively (1) by smelting in reverberatory furnaces
(Tarnowitzeröfen), and (2) by roasting in reverberatory-sintering
furnaces roasted material in the shaft furnace. The factor which
determined whether the treatment was to be effected in the
reverberatory-smelting or in the roasting-sintering furnace was the
percentage of lead and zinc in the ores; those comparatively rich in
lead and poor in zinc being worked up in the former, with partial
production of pig-lead; while those poorer in lead and richer in zinc
were treated in the latter. About two-fifths of the lead ores annually
worked up were charged into the reverberatory-smelting furnaces, and
three-fifths into the sintering furnaces.
In 1900 there were available 10 reverberatory-smelting and nine
sintering furnaces. These were worked exclusively by hand.
The sintered product of the roasting furnaces, and the gray slag from
the reverberatory-smelting furnaces, were transferred to the shaft
furnaces for further treatment, and were therein smelted together with
the requisite fluxes. Eight such furnaces (8 m. high, and 1.4 m., 1.6
m., and 1.8 m. respectively in diameter at the tuyeres), partly with
three and partly with five or eight tuyeres, were at that time in use.
Now that the Huntington-Heberlein process has been completely
installed, the reverberatory-smelting furnaces have been shut down
entirely, and the sintering furnaces also for the most part; all
kinds of lead ore, with a single exception, are worked up by the
Huntington-Heberlein process, irrespective of the contents of lead and
zinc. An exceedingly small proportion of the ore treated, viz., the
low-grade concentrate (Herdschlieche) containing 25 to 35 per cent. Pb,
is still roasted in the old sintering furnace, together with various
between-products (such as dust, fume, scaffoldings, and matte); these
are scorified by the aid of the high percentage of silica in the
material.
For roasting lead ores at the present time there are six round
mechanical roasters of 6 m. diameter, one of 8 m. diameter, and two
ordinary, stationary Huntington-Heberlein furnaces. The latter (which
represent the primitive Huntington-Heberlein furnaces, requiring manual
labor) have recently been shut down, and will probably never be used
again. In the mechanical Huntington-Heberlein furnace, roasting of lead
ore is carried only to such a point that a small portion of the lead
sulphide is converted into sulphate. The desulphurization of the ore
is completed in the so-called converter (made of iron, pear-shaped or
hemispherical in form) in which the charge, up to this stage loosely
mixed, is blown to a solid mass.
Owing to the ready fusibility of this product (which still contains,
as a rule, up to 1.5 per cent. sulphur as sulphide), it is possible to
use shaft furnaces of rather large dimensions; therefore a round shaft
furnace (2.4 m. diameter at the tuyeres, 7 m. high, and furnished with
15 tuyeres) was built. In this furnace nearly the whole of the roasted
ore from the Huntington-Heberlein converters is now smelted, some of
the smaller shaft furnaces being used occasionally. The introduction
of the new process has caused no noteworthy change in the subsequent
treatment of the work-lead.
In the following study I shall discuss the treatment of a given annual
quantity of ore (50,000 tons), which is the actual figure at the
Friedrichshütte at the present time.
1. _Roasting Furnaces._—A reverberatory-smelting furnace used to treat
5 tons of ore in 24 hours; a roasting-sintering furnace, 8 tons.
Assuming the ratios previously stated, the annual treatment by the
former process would be 20,000 tons, and by the latter 30,000 tons.
On the basis of 300 working days per year, and no prolonged stoppages
for furnace repairs (though considering the high temperatures of these
furnaces this record would hardly be expected), there would be required:
20,000 ÷ (5 × 300) = 13.3 (or 13 to 14 reverberatory furnaces).
30,000 ÷ (8 × 300) = 12.5 (or 12 to 13 sintering furnaces).
The capacity of a stationary Huntington-Heberlein furnace is 18 tons;
hence in order to treat the same quantity of ores there would be
required:
50,000 ÷ (18 × 300) = 9.3 (or 9 to 10 Huntington-Heberlein furnaces).
With the revolving-hearth roasters (of 6 m. diameter) working a total
charge of at least 27 tons of ore, there would be required:
50,000 ÷ (27 × 300) = 6.1 (or 6 to 7 roasters).
Still better results are obtained with the 8 m. round roaster, which
has been in operation for some time; in this, 55 tons of ore can be
roasted daily. Three such furnaces would therefore suffice for working
up the whole of the ore charged per annum.
Now, making due provision for reserve furnaces, to work up 50,000 tons
of ore would require:
Reverberatory (15) and sintering furnaces (15) 30
Stationary Huntington-Heberlein furnaces 12
6 m. revolving-hearth furnaces 8
8 m. revolving-hearth furnaces 4
Similar relations hold good regarding the number of workmen
attending the furnaces, there being required, daily, six men for the
reverberatory furnace; eight men for the sintering furnace; ten men for
the stationary; and six men for the mechanical Huntington-Heberlein
furnace; or, for 14 reverberatory furnaces, daily, 84 men; for
sintering furnaces, daily, 104 men; total, 188 men. While for 10
stationary Huntington-Heberlein furnaces, 100 men are required; and
for 7 mechanical Huntington-Heberlein furnaces, daily, 42 men. It is
expected that only 14 men (working in two shifts) will be required to
run the new installation with 8 m. round roasters.
It is true that the exclusion of human labor here has been carried to
an extreme. The roasters and converters will be charged exclusively
by mechanical means; thus every contact of the workmen with the
lead-containing material is avoided until the treatment of the roasted
material in the converters is completed.
From the data given above, the capacity of each individual workman
is readily determined, as follows: With the reverberatory-smelting
furnace, each man daily works up 0.83 tons; with the sintering furnace,
1 ton; with the stationary Huntington-Heberlein furnace, 1.8 tons;
with the 6 m. revolving-hearth furnace, 4.5 tons; and with the 8 m.
revolving-hearth furnace, 11.8 tons.
A significant change has also taken place in coal consumption. Thus,
when working with the reverberatory and sintering furnaces in order to
attain the requisite temperature of 1000 deg. C., there was required
not only a comparatively high-grade coal, but also a large quantity of
it. A reverberatory furnace consumed about 503 kg., a sintering furnace
about 287 kg., of coal per ton of ore. For roasting the ore in the
stationary and also in the mechanical Huntington-Heberlein furnaces, a
lower temperature (at most 700 deg. C.) is sufficient, as the roasting
proper of the ore is effected in the converters, and the sulphur
furnishes the actual fuel. For this reason, the consumption of coal is
much lower. The comparative figures per ton of ore are as follows: In
the reverberatory furnace, 50.3 per cent.; in the sintering furnace,
28.7 per cent.; in the stationary Huntington-Heberlein furnace, 10.3
per cent.; and in the Huntington-Heberlein revolving-hearth furnace,
7.3 per cent.
But there is another technical advantage of the Huntington-Heberlein
process which should be mentioned. It is well known that the
volatilization of lead at high temperatures is an exceedingly
troublesome factor in the running of a lead-smelting plant; the
recovery of the valuable fume is difficult, and requires condensing
apparatus, to say nothing of the unhealthful character of the volatile
lead compounds. This volatilization is of course particularly marked at
the high temperatures employed when working with reverberatory-smelting
furnaces; the same is true, in a somewhat less degree, of the sintering
furnaces. In consequence of the markedly lower temperature to which
the charge is heated in the Huntington-Heberlein furnace, and also of
the peculiar mode of completing the roast in blast-converters, the
production of fume is so reduced that the difference between the values
recovered in the old and the new processes is very striking. Whereas,
in 1900, in working up 12,922 tons of ore in the reverberatory-smelting
furnace, and 14,497 tons in the sintering furnace (27,419 tons in
all), there was recovered 2470 tons (or 9 per cent.) as fume from
the condensers and smoke flues, the quantity of fume recovered, in
1903, fell to 879 tons (or 1.8 per cent.), out of the 48,208 tons of
ore roasted, and this notwithstanding the fact that in the meantime
fume-condensing appliances had been considerably expanded and improved,
whereby the collection was much more efficient.
Lastly, the zinc content of the ores no longer exerts the same
unfavorable influence as in the old process (wherein it was advisable
to subject ore containing much blende to a final washing before
proceeding to the actual metallurgical treatment). In the new process,
the ores are simply roasted without regard to their zinc content. In
this connection it has been found that a considerable proportion of the
zinc passes off with the fume, and that the roasted material usually
contains a quantity of zinc so small that it no longer causes any
trouble in the shaft furnace. It may also be mentioned here that the
ore-dressing plants recently installed in the mines of Upper Silesia
have resulted in a more perfect separation of the blende.
_Shaft Furnaces._—The finished product from the Huntington-Heberlein
blast-converters is of a porous character, and already contains a
part of the flux materials (such as limestone, silica and iron) which
are required for the shaft-furnace charge. It is just these two
characteristics of the roasted product (its porous nature, on the one
hand, leading to its more perfect reduction by the furnace gases; and,
on the other hand, the admixture of fluxes in the molten condition,
resulting in a more complete utilization of the temperature), which,
together with its higher lead and lower zinc content, determine its
ready fusibility. If we further consider that it is possible in the new
process to make the total charge of the shaft furnace richer in lead
than formerly (two-thirds of the total charge as against one-third),
and that a higher blast pressure can be used without danger, it follows
immediately that the capacity of a shaft furnace is much greater by
the new process than by the old method of working. The daily production
of the shaft furnaces on the old and the new process is as shown in the
table given herewith:
─────────────┬─────────────────────────┬─────────┬────────────────────
│ │ CHARGE │ WORK-LEAD
TYPE OF SHAFT│ CHARACTER OF CHARGE │ PER DAY,│ PRODUCED
FURNACE │ │ TONS │ PER DAY, TONS
─────────────┼─────────────────────────┼─────────┼────────────────────
3 tuyeres │{ Gray slag from } │ 36 │ 6 to 7 }
│{ reverberatory } │ │ }
│{ furnaces and } │ │ } Low-
│{ sintered concentrate } │ │ }pressure
│ │ │ } Blast
8 tuyeres │ ” ” │ 36 to 38│ 6 to 8 }
│ │ │ }
3 tuyeres │{ Roasted product of } │ 36 │ 11 to 12 }
│{ Huntington-Heberlein } │ │
│{ process } │ │
│ │ │
8 tuyeres │ ” ” │ 65 to 72│ 24 to 26 } High-
│ │ │ }pressure
15 tuyeres │ ” ” │ 270 │ 90 to 100 } Blast
─────────────┴─────────────────────────┴─────────┴────────────────────
It should be noted that the figure given for the furnace with 15
tuyeres represents the average for 1904; this average is lowered by the
circumstance that during this period there was frequently a deficiency
of roasted material, and the furnace had to work with low-pressure
blast. A truer impression can be gained from the month of March, 1905,
for instance, during which time this furnace worked under normal
conditions; the results are as follows:
The average for March, 1905, was: Ore charged, 8,269.715 tons; coke,
652.441 tons; total, 8,922.156 tons. Or, in 24 hours: Ore charged,
266.765 tons; coke, 21.046 tons; total, 287.811 tons. The production of
work-lead was 3,133.245 tons, or 101.069 tons per day.
The maximum production of roasted ore was 210 tons, on June 30, 1905,
when the total charge was: Ore, 327.38 tons; coke, 25.2 tons; total,
352.58 tons. The quantity of work-lead produced on that day was 120.695
tons, while the largest quantity previously produced in one day was
124.86 tons. It should also be mentioned that the lead tenor of the
slag is almost invariably below 1 per cent.; it usually lies between
0.3 and 0.5 per cent.
As in the case of the roasting furnaces, the productive capacity of
the shaft furnace also comes out clearly if we figure the number
of furnaces required, on the basis of an annual consumption of
50,000 tons of ore. If we consider 1 ton of the roasted material as
equivalent to 1 ton of ore (which is about right in the case of the
Huntington-Heberlein material, but is rather a high estimate in the
case of the product of the sintering furnace), then, in the old process
(where one-third of the charge was lead-bearing material), 12 tons
could be smelted daily. There would therefore be needed at least:
50,000 ÷ (12 × 300) = 14 three-tuyere shaft furnaces.
Since, as already mentioned, the lead-bearing part of the charge
constitutes two-thirds of the whole in the Huntington-Heberlein
process, the number of shaft furnaces of different types, as compared
with the foregoing, would figure out:
3-tuyere shaft furnace, with product of sintering furnace,
50,000 ÷ (12 × 300) = 14 furnaces;
3-tuyere shaft furnace, with product of Huntington-Heberlein furnace,
50,000 ÷ (24 × 300) = 7 furnaces;
8-tuyere shaft furnace, with product of Huntington-Heberlein furnace,
50,000 ÷ (48 × 300) = 3.4 (say 4) furnaces;
15-tuyere shaft furnace, with product of Huntington-Heberlein furnace,
50,000 ÷ (180 × 300) = 1 furnace.
Running regularly and without interruption, the large shaft furnace is
therefore fully capable of coping with the Huntington-Heberlein roasted
material at the present rate of production.
As regards the number of workmen and the product turned out per man,
no such marked difference is produced by the introduction of the
Huntington-Heberlein process in the case of the shaft furnace as there
was noted for the roasting operation. This is chiefly due to the fact
that the work which requires the more power (such as charging of the
furnaces, conveying away the slag and pouring the lead) can be executed
only in part by mechanical means. Nevertheless, it will be seen from
the table given herewith that, on the one hand, the number of men
required for the charge worked up is smaller; and, on the other, the
product turned out per man has risen somewhat.
─────────┬─────────┬────────┬──────────┬────────┬─────────────┬───────
TYPE OF │CHARACTER│ CHARGE │NUMBER OF │ CHARGE │DAILY OUTPUT │OUTPUT
SHAFT │OF CHARGE│PER DAY,│FURNACEMEN│PER MAN,│OF WORK-LEAD,│PER MAN,
FURNACE │ │ TONS │ │ TONS │ TONS │ TONS
─────────┼─────────┼────────┼──────────┼────────┼─────────────┼───────
3 tuyere│ A │ 36 │ 6 │ 6.0 │ 6 │ 1.0
8 tuyere│ B │ 38 │ 6 │ 6.3 │ 8 │ 1.3
3 tuyere│ C │ 36 │ 6 │ 6.0 │ 12 │ 2.0
8 tuyere│ D │ 72 │ 12 │ 6.0 │ 26 │ 2.1
15 tuyere│ E │ 270 │ 34 │ 7.9 │ 90 │ 2.6
─────────┴─────────┴────────┴──────────┴────────┴─────────────┴───────
┌──────────┬──────────────────────────────────────┐
│ CHARACTER│ CHARACTER OF CHARGE │
│ OF CHARGE│ │
│ CODE │ │
├──────────┼──────────────────────────────────────┤
│ A │ Sintered concentrate and gray slag │
│ B │ from reverberatory furnace. │
│ B │ Gray slag from reverberatory furnace.│
│ C │ Huntington-Heberlein product. │
│ D │ Huntington-Heberlein product. │
│ E │ Huntington-Heberlein product. │
└──────────┴──────────────────────────────────────┘
A slight difference only is produced by the new process in the
consumption of coke; the economy is a little over 1 per cent., the
coke consumed being reduced from 9.39 per cent. to 8.17 per cent. of
the total charge. But with the high price of coke, even this small
difference represents a considerable lowering of the cost of production.
With the great increase in the blast pressure, it would be supposed
that the losses in fume would be much greater than with the former
method of working. But this is not the case; on the contrary, all
experience so far shows that there is much less fume developed. In
1904, for instance, the shaft-furnace fume recovered in the condensing
system amounted to only 1.06 per cent. of the roasted material, or
0.64 per cent. of the total charge, as against 2.03 and 1.0 per cent.,
respectively, in former years. The observations made on the quantity of
flue dust carried away with the gases escaping into the air through the
stack showed that it is almost nil.
Now, from the loss in fume being slight, from the tenor of lead in the
slag being low, and, on the one hand, from the quantity of lead-matte
produced being much less than before, while on the other the losses in
roasting the ore are greatly reduced—from all these considerations, it
is clear that the total yield must have been much improved. As a matter
of fact, the yield of lead and silver has been increased by at least 6
to 8 per cent.
_Economic Results._—As regards the economical value of the new process,
for obvious reasons no data can be furnished of the exact expenditure,
i.e., the actual total cost of roasting and smelting the ore. But
this at least is placed beyond doubt by what has been developed above,
namely, that considerable saving must be effected in the roasting,
and especially in the smelting, as compared with the former mode of
working. If we take into account only the economy which is gained
in wages through the increase in the material which one workman can
handle, and that resulting from the reduced consumption of coal and
coke, these alone will show sufficiently that an important diminution
of working cost has taken place. The objection which might be raised,
that the saving effected by reducing manual labor may be neutralized
by the expense of mechanical power (actuating the roasters, furnishing
the compressed blast, etc.), cannot be regarded as justified, as the
cost of mechanical work is comparatively low. Thus, for instance, the
large 8 m. furnace and the small, round furnaces require 15 h.p. if
worked by electricity. According to an exact calculation, the cost
(to the producer) of the h.p. hour, inclusive of machinery, figures
out to 3.6 pfennigs (0.9c.); hence the daily expense for running the
revolving-hearth furnaces amounts to: 15 × 3.6 pfg. × 24 = 12.96 marks
($3.42). As the seven furnaces together work up: (6 × 27) + 55 = 217
tons of ore, the cost per ton of ore is about 0.06 mark (1.5c.).
The requisite blast is produced by means of single-compression Encke
blowers, of which one is quite sufficient when running at full load,
and then consumes 34 h.p. The daily expenses are accordingly: 34 × 3.6
pfg. × 24 = 29.28 marks ($7.32); or per ton of ore, 29.28 ÷ 217 = 0.14
mark (3.5c.). Therefore the total expense for the mechanical work in
roasting the ore amounts to 0.06 + 0.14 = 0.20 mark (5c.).
However, the cost of roasting is much more affected by the expense
for keeping the furnaces in repair; another important factor is the
acquisition and maintenance of the tools. Both in the case of the
sintering and also the reverberatory-smelting furnace, the cost of
keeping in repair was high; the consumption of iron was especially
large, owing to the rapid wear of the tools. This was not surprising,
considering that a notably higher temperature prevailed in the
reverberatory and sintering furnaces than in the new roasters, in which
the temperature strictly ought not to rise above 700 deg. C. But in the
old type of furnace the high temperature and the constant working with
the iron tools caused their rapid wear, thus creating a large item for
iron and steel and smith work. In the new process (and more especially
in the revolving-hearth roasters) this disadvantage does not arise. In
this case there is practically no work on the furnace, and the wear
and tear of iron is small. Also, the cost of keeping the furnaces
in repair when working regularly is small as compared with the old
process. In the year 1900, for instance, the cost of maintenance and
tools for the reverberatory and sintering furnaces came to 20,701.93
marks ($5,175.48) for treating 27,419.75 tons of ore. Per ton of ore,
this represents 0.75 mark (19c.). In the year 1903, on the other
hand, only 9,074.17 marks ($2,268.54) were expended, although 48,208
tons of ore were worked up in the three stationary and six mechanical
Huntington-Heberlein furnaces. The cost of maintenance was, therefore,
in this case 0.18 mark (4.5c.) per ton of ore.
In the cost of smelting in the shaft furnace, only a slight difference
in favor of the Huntington-Heberlein process is found if the estimate
is based on the total charge; but a marked difference is shown if it is
referred to the lead-bearing portion of the charge, or to the work-lead
produced. Thus the cost of maintenance and total cost of smelting,
figured for one ton of ore, without taking into account general
expenses, have been tabulated as follows:
────────────────────────────┬────────────────────────────────
│REDUCTION IN EXPENSES PER TON OF
├────────┬──────────┬────────────
│ TOTAL │ LEAD ORE │ WORK-LEAD
│ CHARGE │ │
────────────────────────────┼────────┼──────────┼────────────
(_a_) Cost of maintenance │ 0.01M │ 0.38M │ 0.67M
│(0.25c) │ (9.5c) │ (16.75c)
│ │ │
(_b_) Total cost of smelting│ 0.20M │ 6.46M │ 11.48M
│ (5c) │ ($1.615) │ ($2.87)
────────────────────────────┴────────┴──────────┴────────────
The marked reduction in the expenses, as referred to the lead-ore and
the work-lead produced, is determined (as was pointed out above) by the
greater lead content of the charge, and by the larger yield of lead
consequent thereon. The advantage of longer smelting campaigns (which
ultimately were mostly prolonged to one year) also makes itself felt;
it would be still more marked, if the shaft furnace (which was still in
working condition after it was blown out) had been run on for some time
longer.
Finally, if we examine the question of the space taken up by the plant
(which, owing to the scarcity of suitably located building sites,
would have been important at the Friedrichshütte at the time when the
quantity of ore treated was suddenly doubled), here again we shall
recognize the great advantage which this establishment has gained from
the Huntington-Heberlein process.
As was calculated above, there would have been required 15
reverberatory and 15 sintering furnaces to cope with the quantity of
ore treated. As a reverberatory requires, in round numbers, 120 sq. m.
(1290 sq. ft.), and a sintering furnace 200 sq. m. (2153 sq. ft.); and
as fully 100 sq. m. (1080 sq. ft.) must be allowed for each furnace for
a dumping ground, therefore the 15 reverberatory furnaces would have
required an area of 15 × 120 + 15 × 100 = 3300 sq. m.; the 15 sintering
furnaces would have required 15 × 200 + 15 × 100 = 4500 sq. m.; in
all 3300 + 4500 = 7800 sq. m. (83,960 sq. ft.). The 12 stationary
Huntington-Heberlein furnaces (built together two and two) would take
up a space of 6 × 200 + 12 × 100 = 2400 sq. m. (25,830 sq. ft.).
Similarly, 8 small furnaces would require 8 × 100 + 8 × 100 = 1600 sq.
m. (17,222 sq. ft.); while for the new installation of four 8-meter
revolving-hearth furnaces and 10 large converters, only 1320 sq. m.
(14,120 sq. ft.) have been allowed.
For shaft furnaces with three or eight tuyeres, which were run with
low-pressure blast for the material roasted on the old plan, the total
area built upon was 18 × 16.5 = 297 sq. m.; while a further area of 18
× 14 = 250 sq. m. was hitherto provided, and was found sufficient for
dumping slag when working regularly. Therefore, the installation of
shaft furnaces formerly in existence, after requisite enlargement to
14 furnaces, would have demanded a space of 7 × 297 + 7 × 250 = 3829
sq. m. (42,215 sq. ft.). If four of the small shaft furnaces had been
reconstructed for eight tuyeres, and run with Huntington-Heberlein
roasted material, using high-pressure blast, the area occupied would
have been reduced to 2 × 297 + 2 × 250 sq. m. = 1094 sq. m. (11,776 sq.
ft.).
Still more favorable are the conditions of area required in the case of
the large shaft furnace. This furnace stands in a building covering an
area of 350 sq. m. (3767 sq. ft.), which is more than sufficient room.
The slag-yard (situated in front of this building, and amply large
enough for 36 hours’ run) has an area of 250 sq. m. (2691 sq. ft.);
thus the space occupied by the large shaft furnace, including a yard of
170 sq. m. (1830 sq. ft.), is in all 780 sq. m. (8396 sq. ft.).
After completion of the new roasting plant and the large shaft furnace
in connection with it, there would be occupied 1320 + 780 = 2100 sq.
m. (2260 sq. ft.); and if the system of reverberatory and sintering
furnaces had been continued (with the requisite additions thereto and
to the old shaft-furnace system), there would have been required 11,629
sq. m. (125,214 sq. ft.). In the estimate above given no regard has
been paid to any of the auxiliary installations (dust chambers, etc.),
which, just as in the case of the old process, would have had to be
provided on a large scale.
It is of course self-evident that both the principal and the auxiliary
installations in the old process would not only have involved a high
first cost, but would also, on account of their extensive dimensions,
have caused considerably greater annual expense for maintenance.
THE HUNTINGTON-HEBERLEIN PROCESS FROM THE HYGIENIC STANDPOINT[30]
BY A. BIERNBAUM
(October 14, 1905)
With regard to the hygienic improvements which the Huntington-Heberlein
process offers, we must first deal with the questions: What were
the sources of danger in the old process, and in what way are these
now diminished or eliminated? The only danger which enters into
consideration is lead-poisoning, other influences detrimental to health
being the same in one process as the other.
With the reverberatory-smelting and roasting-sintering furnaces, the
chief danger of lead-poisoning lies in the metallic vapor evolved
during the withdrawal of the roasted charge from the furnace. It is
true that appliances may be provided, by which these vapors are drawn
off or led back into the furnace during this operation; but, even
working with utmost care, it is impossible to insure the complete
elimination of lead fumes, especially in wheeling away the pots
filled with the red-hot sintered product. Moreover, the work at the
reverberatory-smelting and roasting-sintering furnaces involves great
physical exertion, wherefore the respiratory organs of the workmen
are stimulated to full activity, while the exposure to the intense
heat causes the men to perspire freely. Hence, as has been established
medically, the absorption of the poisonous metallic compounds (which
are partially soluble in the perspiration) into the system is favored
both by inhalation of the lead vapor and by its penetration into the
pores of the skin, opened by the perspiration.
A further danger of lead-poisoning was occasioned by the frequently
recurring work of clearing out the dust flues. The smoke from the
reverberatory-smelting furnace especially contained oxidized lead
compounds, which on absorption into the human body might readily be
dissolved by the acids of the stomach, and thus endanger the health of
the workmen.
In the Huntington-Heberlein furnaces, on the other hand, although the
charge is raked forward and turned over by hand, it is not withdrawn,
as in the old furnaces, by an opening situated next to the fire, but
is emptied at a point opposite into the converters which are placed
in front of the furnace. Moreover, the converters are filled with the
charge at a much lower temperature. Inasmuch as this charge has already
cooled down considerably, there can be practically no volatilization of
lead. The small quantity of gas which may nevertheless be evolved is
drawn off by fans through hoods placed above the converters.
A further improvement, from the hygienic point of view, is in the use
of the mechanical furnaces, from which the converters can be filled
automatically (almost without manual labor, and with absolute exclusion
of smoke). The converters are then placed on their stands and blown.
This work also is carried out under hoods, as gas-tight as possible,
furnished with a few closable working apertures. During the blowing
of the material, the work of the attendant consists solely in keeping
up the charge by adding more cold material and filling any holes that
may be formed. It does not entail nearly as much physical strain as
the handling of the heavy iron tools and the continued exposure of the
workmen to the hottest part of the furnace, which the former roasting
process involved.
Some experiments carried out with larger converters (of 4 and 10
ton capacity) have indicated the direction in which the advantages
mentioned above may probably be developed to such a point that the
danger of lead-poisoning need hardly enter into consideration. Both
the charging of the revolving-hearth furnaces and the filling of the
converters are to be effected mechanically. Furthermore, in the case
of the large converters the filling up of holes becomes unnecessary,
and no manual work of any kind is required during the whole time
of blowing. The converters can be so perfectly enclosed in hoods
that the escape of gases into the working-rooms becomes impossible,
and lead-poisoning of the men can occur only under quite unusual
circumstances.
The beneficial influence on the health of the workmen attending
on the roasting furnaces, occasioned by the introduction of the
Huntington-Heberlein process, can be seen from the statistics of
sickness from lead-poisoning for the years 1902 to 1904, as given
herewith:
─────────┬──────┬──────┬──────────────────────────────┬───────────────
│ │ │ LEAD-POISONING │ CASES
│ │ ├─────────────┬────────────────┤ CONTRACTED
│ │ │NO. OF CASES │DAYS OF SICKNESS│AT REVER.│ AT
─────────┼──────┼──────┼─────┬───────┼───────┬────────┤ AND │H. H. SICKNESS
METHOD OF│ YEAR │NO. OF│TOTAL│PER 100│ TOTAL │PER 100 │ SINT. │ FUR.
WORKING │ │ MEN │ │PERSONS│ │PERSONS │ FUR. │
─────────┼──────┼──────┼─────┼───────┼───────┼────────┼─────────┼─────
│ │ │ │ │ │ │ │
Old │{ 1902│ 93 │ 15 │ 16.1 │ 246 │ 264.5 │ 11 │ 4
│{ 1903│ 86 │ 12 │ 13.9 │ 222 │ 258.1 │ 7 │ 5
│ │ │ │ │ │ │ │
H.-H. │ 1904│ 87 │ 8 │ 9.2 │ 242 │ 278.2 │ 6 │ 2
─────────┴──────┴──────┴─────┴───────┴───────┴────────┴─────────┴─────
This shows a gratifying decrease in the number of cases, namely, from
16.1 to 9.2 per cent.; this decrease would have been still greater if
Huntington-Heberlein furnaces had been in use exclusively. However,
most of the time two or three sintering furnaces were fired for
working up by-products, 16 to 18 men being engaged on that work. The
Huntington-Heberlein furnaces alone (at which, in the year 1904, 69 men
in all were occupied) show only 2.9 per cent. of cases. That the number
of days of illness was not reduced is due to the fact that the cases
among the gang of men working at the sintering furnaces were mostly of
long standing and took some time to cure.
The noxious effects upon the health of the workmen in running the
shaft furnaces are due to the fumes from the products made in this
operation, such as work-lead, matte and slag, which flow out of the
furnace at a temperature far above their melting points. Even with
the old method of running the shaft furnaces the endeavor has always
been to provide as efficiently as possible against the danger caused
by this volatilization, and, wherever feasible, to install safety
appliances to prevent the escape of lead vapors into the work-rooms;
but these measures could not be made as thorough as in the case of the
Huntington-Heberlein process.
The principal work in running the shaft furnaces, aside from the
charging, consists in tapping the slag and pouring out the work-lead.
Other unpleasant jobs are the barring down (which in the old process
had to be done frequently) and the cleaning out of the furnace after
blowing out.
In the old process the slag formed in the furnace flows out
continuously through the tap-hole into iron pots placed in front of
the spout. A number of such pots are so arranged on a revolving table
that as soon as one is filled the next empty can be brought up to the
duct; thus the slag first poured in has time to cease fuming and to
solidify before it is removed. The vapors arising from the slag as it
flows out are conveyed away through hoods. At the same time with the
slag, lead matte also issues from the furnace. Now the greater the
quantity of lead matte, the more smoke is also produced; and, with
the comparatively high proportion of lead matte resulting from the
old process, the quantity of smoke was so great that the ventilation
appliances were no longer sufficient to cope with it, thus allowing
vapors to escape into the work-room.
The work-lead collects at the back of the furnace in a well, from which
it is from time to time ladled into molds placed near by. If the lead
is allowed to cool sufficiently in the well, it does not fume much in
the ladling out. But when the furnace runs very hot (which sometimes
happens), the lead also is hotter and is more inclined to volatilize.
In this event the danger of lead-poisoning is very great, for the
workman has to stand near the lead sump.
A still greater danger attends the work of barring down and cleaning
out the furnace. The barring down serves the purpose of loosening
the charge in the zone of fusion; at the same time it removes any
crusts formed on the sides of the furnace, or obstructions stopping
up the tuyeres. With the old furnaces, and their strong tendency to
crust, this work had to be undertaken almost every day, the men being
compelled to work for rather a long time and often very laboriously
with the heavy iron tools in the immediate neighborhood of the glowing
charge, the front of the furnace being torn open for this purpose. In
this operation they were exposed without protection to the metallic
vapors issuing from the furnace, inasmuch as the ventilating appliances
had to be partially removed during this time, in order to render it at
all possible to do the work.
In a similar manner, but only at the time of shutting down a shaft
furnace, the cleaning out (that is to say, the withdrawing of no
longer fused but still red-hot portions of the charge left in the
furnace) is carried out. In this process, however, the glowing material
brought out could be quenched with cold water to such a point that the
evolution of metallic vapors could be largely avoided.
Lastly, the mode of charging of the shaft furnace is also to be
regarded as a cause of poisoning, inasmuch as it is impossible to
avoid entirely the raising of dust in the repeated act of dumping and
turning over the materials for smelting, in preparing the mix, and in
subsequently charging the furnace.
By the introduction of the Huntington-Heberlein process, all these
disadvantages, both in the roasting operation and in running the shaft
furnaces, are in part removed altogether, in part reduced to such a
degree that the danger of injury is brought to a minimum.
In furnaces in which the product of the Huntington-Heberlein roast
is smelted, the slag is tapped only periodically at considerable
intervals; and, as there is less lead matte produced than formerly, the
quantity of smoke is never so great that the ventilating fan cannot
easily take care of it. There is therefore little chance of any smoke
escaping into the working-room.
As the production of work-lead, especially in the case of the large
shaft furnace, is very considerable, so that the lead continually
flows out in a big stream into the well, the hand ladling has to
be abandoned. Therefore the lead is conducted to a large reservoir
standing near the sump, and is there allowed to cool below its
volatilizing temperature. As soon as this tank is full, the lead is
tapped off and (by the aid of a swinging gutter) is cast into molds
ready for this purpose. Both the sump and the reservoir-tank are placed
under a fume-hood. The swinging gutter is covered with sheet-iron lids
while tapping, so that any lead volatilized is conveyed by the gutter
itself to a hood attached to the reservoir; thus the escape of metallic
vapors into the working space is avoided, as far as possible.
This method of pouring does not entail the same bodily exertion as the
ladling of the lead; moreover, as it requires but little time, it gives
the workmen frequent opportunity to rest.
But one of the chief advantages of the Huntington-Heberlein process
lies in the entire omission of the barring down. If the running of the
shaft furnace is conducted with any degree of care, disorders in the
working of the furnace do not occur, and one can rely on a perfectly
regular course of the smelting process day after day. No formation
of any crusts interfering with the operation of the furnace has been
recorded during any of the campaigns, which have, in each case, lasted
nearly a year.
As regards the cleaning out of the furnace, this cannot be avoided
on blowing out the Huntington-Heberlein shaft furnace; but at most
it occurs only once a year, and can be done with less danger to the
workmen, owing to the better equipment.
Further, the charge is thrown straight into the furnace (in the case
of the large shaft furnace); thus the repeated turning over of the
smelting material, as formerly practised, becomes unnecessary, and the
deleterious influence of the unavoidable formation of dust is much
diminished.
The accompanying statistics of sickness due to lead-poisoning in
connection with the operation of the shaft furnace (referring to the
same period of time as those given above for the roasting furnaces)
confirm the above statements.
────┬──────────┬────────────────────────────────────────────
│ │ LEAD-POISONING—SHAFT FURNACES
│ ├─────────────────────┬──────────────────────
YEAR│NO. OF MEN│ CASES │ DAYS OF ILLNESS
│ ├─────┬───────────────┼─────┬────────────────
│ │TOTAL│PER 100 PERSONS│TOTAL│PER 100 PERSONS
────┼──────────┼─────┼───────────────┼─────┼────────────────
1902│ 250 │ 58 │ 23.2 │ 956 │ 382.4
1903│ 267 │ 59 │ 22.1 │1044 │ 391.0
1904│ 232 │ 24 │ 10.3 │ 530 │ 228.4
────┴──────────┴─────┴───────────────┴─────┴────────────────
If it were possible to make the necessary distinctions in the case of
the large shaft furnace, the diminution in sickness from lead-poisoning
would be still more apparent; for, among the furnace attendants proper,
there has been no illness; all cases of poisoning have occurred among
the men who prepare the charge, who break up the roasted material, and
others who are occupied with subsidiary work. Some of these are exposed
to illness through their own fault, owing to want of cleanliness, or to
neglect of every precautionary measure against lead-poisoning.
Thus far we have dealt only with the advantages and improvements of the
Huntington-Heberlein process; we will now, in conclusion, consider also
its disadvantages.
The chief drawback of the new process lies in the difficulty of
breaking up the blocks of the roasted product from the converters, a
labor which, apart from the great expense involved, is also unhealthy
for the workmen engaged thereon. Seemingly this evil is still further
increased by working with larger charges in the 10 ton converters, as
projected; but in this case it is proposed to place the converters in
an elevated position, and to cause the blocks to be shattered by their
fall from a certain hight, so that further breaking up will require
but little work. Trials made in this direction have already yielded
satisfactory results, and seem to promise that the disadvantage will in
time become less important.
Another unpleasant feature is the presence (in the waste gases from the
converters) of a higher percentage of sulphur dioxide, the suppression
of which, if it is feasible at all, might be fraught with trouble and
expense.
That the roaster gases from the reverberatory-smelting and sintering
furnaces did not show such a high percentage of sulphur dioxide must
be ascribed chiefly to the circumstance that the roasting was much
slower, and that the gases were largely diluted with air already at the
point where they are formed, as the work must always be done with the
working-doors open. In the Huntington-Heberlein process, on the other
hand, the aim is to prevent, as far as possible, the access of air from
outside while blowing the charge. The more perfectly this is effected,
and the greater the quantity of ore to be blown in the converters, the
higher will also be the percentage of sulphur dioxide in the waste
gases. This circumstance has not only furnished the inducement, but it
has rendered it possible to approach the plan of utilizing the sulphur
dioxide for the manufacture of sulphuric acid. If this should be done
successfully (which, according to the experiments carried out, there
is reasonable ground to expect), the present disadvantage might be
turned into an advantage. This has the more significance because an
essential constituent of the lead ore—the sulphur—will then no longer,
as hitherto, have to be regarded as wholly lost.[31]
THE HUNTINGTON-HEBERLEIN PROCESS
BY THOMAS HUNTINGTON AND FERDINAND HEBERLEIN
(May 26, 1906)
This process for roasting lead sulphide ores has now fairly
established itself in all parts of the world, and is recognized by
metallurgical engineers as a successful new departure in the method of
desulphurization. It offers the great advantage over previous methods
of being a more scientific application of the roasting reactions (of
the old well-used formulæ PbS + 3O = PbO + SO₂ and PbS + PbSO₄
+ 2O = 2PbO + 2SO₂) and admits of larger quantities being handled
at a time, so that the use of fuel and labor are in proportion to the
results achieved, and also there is less waste all around in so far
as the factors necessary for the operation—fuel, labor and air—can
be more economically used. The workman’s time and strength are not
employed in laboriously shifting the ore from one part of the furnace
to another with a maximum amount of exertion and a minimum amount of
oxidation. The fuel consumed acts more directly upon the ore during the
first part of the process in the furnace and its place is taken by the
sulphur itself during the final and blowing stage, so that during the
whole series of operations more concentrated gases are produced and
consequently the large excess of heated air of the old processes is
avoided to such an extent that the gases can be used for the production
of sulphuric acid.
With a modern well-constructed plant practically all the evils of
the old hand-roasting furnaces are avoided, and besides the notable
economy achieved by the H.-H. process itself, the health and well-being
of the workmen employed are greatly advanced, so that where hygienic
statistics are kept it is proved that lead-poisoning has greatly
diminished. It is only natural, therefore, that the H.-H. process
should have been a success from the start, popular alike with managers
and workmen once the difficulties inseparable from the introduction of
any new process were overcome.
Simple as the process now appears, however, it is the result of many
years of study and experiment, not devoid of disappointments and at
times appearing to present a problem incapable of solution. The first
trials were made in the smelting works at Pertusola, Italy, as far
back as 1889, where considerable sums were devoted every year to this
experimental work and lead ore roasting was almost continuously on the
list of new work from 1875 on.
It may be interesting to mention that at this time the Montevecchio
ores (containing about 70 per cent. lead and about 15 per cent.
sulphur, together with a certain amount of zinc and iron) were
considered highly refractory to roast, and the only ores approved of
by the management of the works at this date were the Monteponi and
San Giovanni first-class ores (containing about 80 per cent. lead),
and the second-class carbonates (with at least 60 per cent. lead and
5 per cent. sulphur). It must be noted that a modified Flintshire
reverberatory process was in use in the works, which could deal
satisfactorily only with this class of ore, so that, as these easy ores
diminished in quantity every year and their place was taken by the
“refractory” Montevecchio type, the roasting problem was always well to
the front at the Pertusola works.
It may be asserted that almost every known method of desulphurization
was examined and experimented upon on a large scale. Gas firing was
exclusively used on certain classes of ores for several years with
considerable success, and revolving furnaces of the Brückner type—gas
fired—were also tried. Although varying degrees of success were
obtained, no really great progress was made in actual desulphurization;
methods were cheapened and larger quantities handled at a time, but
the final product—whether sintered or in a pulverulent state—seldom
averaged much under 5 per cent. sulphur, while the days of the
old “gray slags” (1 per cent. to 2 per cent. sulphur) from the
reverberatories totally disappeared, together with the class of ores
which produced them.
During the long period of these experiments in desulphurization various
facts were established:
(1) That sulphide of lead—especially in a pulverulent state—could not
be desulphurized in the same way as other sulphides, such as sulphides
of iron, copper, zinc, etc., because if roasted in a mechanical
furnace the temperature had to be kept low enough to avoid premature
sintering, which would choke the stirrers and cause trouble by the
ore clogging on the sides and bottom of the furnace. If, however, the
ore was roasted in a “dry state” at low temperature, a great deal of
sulphur remained in the product as sulphate of lead, which was as
bad for the subsequent blast-furnace work as the sulphide of lead
itself. When air was pressed through molten galena—in the same way as
through molten copper matte—a very heavy volatilization of lead took
place, while portions of it were reduced to metal or were contained as
sulphide in the molten matte, so that a good product was not obtained.
(2) That no complete dead roast of lead ores could be obtained unless
the final product was thoroughly smelted and agglomerated.
(3) That a well roasted lead ore could be obtained by oxidizing the PbS
with compressed air, after the ore had been suitably prepared.
(4) That metal losses were mainly due to the excessive heat produced in
the oxidation of PbS to PbO, and other sulphides present in the ore.
It was by making use of these facts that the H.-H. roasting process
was finally evolved, and by carefully applying its principles it is
possible to desulphurize completely the ore to a practically dead roast
of under 1 per cent. sulphur; in practice, however, such perfection
is unnecessary and a well agglomerated product with from 2 to 2.5 per
cent. sulphur is all that is required. During some trials in Australia,
where a great degree of perfection was aimed at, a block of over 2000
tons of agglomerated, roasted ore was produced containing 1 per cent.
sulphur (as sulphide); as the ores contained an average of about 10 per
cent. Zn, this was a very fine result from a desulphurization point
of view, but it was not found that this 1 per cent. product gave any
better results in the subsequent smelting in the blast furnace than
later on a less carefully prepared material containing 2.5 per cent.
sulphur.
In the early stages of experiment the great difficulty was to obtain
agglomeration without first fusing the sulphides in the ore, and
turning out a half-roasted product full of leady matte. Simple as the
thing now is, it seemed at times impossible to avoid this defect, and
it was only by a careful study of the effects of an addition of lime,
Fe₂O₃ or Mn₂O₃, and their properties that the right path
was struck. Before the introduction of the H.-H. process lime was
only used in the reverberatory process (Flintshire and Tarnowitz) to
stiffen the charge, but as Percy tells us that after its addition the
charge was glowing, it must have had a chemical as well as a mechanical
effect. In recognition of this fact fine caustic lime or crushed
limestone was mixed with the ore _before_ charging it into the furnace
and exposing it to an oxidizing heat.
It was thought probable that a dioxide of lime might be temporarily
formed, which in contact with PbS would be decomposed immediately after
its formation, or that the CaO served as _Contactsubstanz_ in the same
way as spongy platinum, metallic silver, or oxide of iron. As CaSO₄
and not CaSO₃ is always found in the roasted ore, this may prove
that CaO is really a contact substance for oxygen (see W. M. Hutchings,
_Engineering and Mining Journal_, Oct. 21, 1905, Vol. LXXX, p. 726).
The fact that the process works equally well with Fe₂O₃ instead
of CaO speaks against the theory of plumbate of lime. Whatever theory
may be correct, the fact remains that CaO assists the roasting process
and that by its use the premature agglomeration of the sulphide ore is
avoided. A further advantage of lime is that it keeps the charge more
porous and thus facilitates the passage of the air.
The shape and size of the blowing apparatus best adapted for the
purpose in view occupied many months; starting from very shallow
pans or rectangular boxes several feet square with a few inches of
material over a perforated plate, it gradually resolved itself into the
cone-shaped receptacle—holding about a ton of ore—as first introduced
together with the process. In later years and in treating larger
quantities a more hemispherical form has been adopted, containing up to
15 tons of ore.
It is probable about eight years were employed in actually working out
the process before it was introduced on any large scale at Pertusola,
but by the end of 1898 the greater part of the Pertusola ores were
treated by the process. Its first introduction to any other works was
in 1900, when it was started outside its home for the first time at
Braubach (Germany). Since then its application has gradually extended,
proceeding from Europe to Australia and Mexico and finally to America
and Canada, where recognition of its merits was more tardy than
elsewhere. It is now practically in general use all over the world and
is recognized as a sound addition to metallurgical progress. It is
doubtless only a step in the right direction and with its general use
a better knowledge of its principles will prevail, so that its future
development in one direction or another, as compared with present
results, may show some further progress.
The present working of the H.-H. process still follows practically the
original lines laid down, and by preliminary roasting in a furnace
with lime, oxide of iron, or manganese (if not already contained in
the ore), prepares the ore for blowing in the converter. Mechanical
furnaces have been introduced to the entire exclusion of the old
hand-roasters, and the size of the converters has been gradually
increased from the original one-ton apparatus successively to 5, 7
and 10 ton converters; at present some for 15 tons are being built in
Germany and will doubtless lead to a further economy.
The mechanical furnace at present most in use is a single-hearth
revolving furnace with fixed rabbles, the latest being built with a
diameter of 26½ ft. and a relatively high arch to ensure a clear flame
and rapid oxidation of the ore. The capacity of these furnaces varies,
of course, with the nature of the ores to be treated, but with ordinary
lead ores (European and Australian practice) of from 50 per cent. to 60
per cent. lead and 14 per cent, to 18 per cent. sulphur, the average
capacity may be taken at about 50 to 60 tons of crude ore per day of
24 hours. The consumption of coal with a well-constructed furnace is
very low and is always under 8 per cent.—6 per cent. being perhaps the
average. These furnaces require very little attention, being automatic
in their charging and discharging arrangements.
The ore on leaving the furnace is charged into the converters by
various mechanical means (Jacob’s ladders, conveyors, etc.). The
converter charge usually consists of some hot ore direct from the
furnace, on top of which ore is placed which has been cooled down by
storage in bins or by the addition of water. The converter is generally
filled in two charges of five tons each, and the blowing time should
not be more than 4 to 6 hours. The product obtained should be porous
and well agglomerated, but easily broken up, tough melted material
being due to an excess of silica and too much lead sulphide. Attention,
therefore, to these two points (good preliminary roasting and
correction of the charge by lime) obviates this trouble. This roasted
ore should not contain more than about 1.5 to 2 per cent. sulphur,
and in a modern blast furnace gives surprisingly good results, the
matte-fall being in most cases reduced to nothing, and the capacity of
the furnace is largely increased, while the slags are poorer.
If the converter charge has been properly prepared, the blowing
operation proceeds with the greatest smoothness and requires very
little attention on the part of the workmen, the heat and oxidation
rise gradually from the bottom and volatilization losses remain low, so
that it is possible, if desired, to produce hot concentrated sulphurous
gases suitable for the manufacture of sulphuric acid.
Besides the actual economy obtained in roasting ores by the process,
a great feature of its success has been the remarkable improvement
in smelting and reducing the roasted ore as compared with previous
experience. This is due to the nature of the roasted material, which,
besides being much poorer in sulphur than was formerly the case, is
thoroughly porous and well agglomerated and contains—if the original
mixture is properly made—all the necessary slagging materials itself,
so that it practically becomes a case of smelting slags instead of ore,
and to an expert the difference is evident.
Experience has shown that on an average the improvement in the capacity
of the blast furnace may be taken at about 50 to 100 per cent., so that
in works using the H.-H. process—after its complete introduction—about
half the blast furnaces formerly necessary for the same tonnage were
blown out. The matte-fall has become a thing of the past, so that,
except in those cases where some matte is required to collect the
copper contained in the ores, lead matte has disappeared and the
quantity of flue dust as well as the lead and silver losses have been
greatly reduced.
Referring to the latest history of the H.-H. process, and the theory
of direct blowing, it may be remarked—putting aside all legal
questions—that the idea, metallurgically speaking, is attractive, as it
would seem that by eliminating one-half of the process and blowing the
ores direct without the expense of a preliminary roast a considerable
economy should be effected. Upon examination, however, this supposed
economy and simplicity is not at all of such great importance, and
in many cases, without doubt, would be retrogressive in lead ore
smelting rather than progressive. When costs of roasting in a furnace
are reduced to such a low figure as can be obtained by using 50 ton
furnaces and 10 to 15 ton converters, there is very little margin
for improvement in this direction. From the published accounts of
the Tarnowitz smelting works (the _Engineering and Mining Journal_,
Sept. 23, 1905, Vol. LXXX, p. 535) the cost of mechanical preliminary
roasting cannot exceed 25c. per ton, so that even assuming direct
blowing were as cheap as blowing a properly prepared material, the
total economy would only be the above figure, viz., 25c.; but this is
far from being the case.
Direct blowing of a crude ore is considerably more expensive than
dealing with the H.-H. product, because of necessity the blowing
operation must be carried out slowly and with great care so as to avoid
heavy metal losses, and whereas a pre-roasted ore can be easily blown
in four hours and one man can attend to two or three 10 ton converters,
the direct blowing operation takes from 12 to 18 hours and requires the
continual attention of one man. In the first case the cost of labor
would be: One man at say $3 for 50 tons (at least), i.e., 6c. per
ton, and in the second case one man at $3 for 10 tons (at the best),
i.e., 30c., a difference in favor of pre-roasting of 24c., so that any
possible economy would disappear. Furthermore, as the danger of blowing
upon crude sulphides for 12 or 18 hours is greater as regards metal
losses than a quick operation of four hours, it is very likely that
instead of an economy there would be an increase in cost, owing to a
greater volatilization of metals.
These remarks refer to ordinary lead ores with say 50 per cent. lead
and about 14 per cent. sulphur. With ores, however, such as are
generally treated in the United States the advantages of pre-roasting
are still more evident. These ores contain about 10 to 15 per cent.
lead, 30 to 40 per cent. sulphur, 20 to 30 per cent. iron, 10 per cent.
zinc, 5 per cent. silica, and lose the greater part of the pyritic
sulphur in the preliminary roasting, leaving the iron in the form of
oxide, which in the subsequent blowing operation acts in the same
way as lime. For this reason the addition of extra fluxes, such as
limestone, gypsum, etc., to the original ore is not necessary and only
a useless expense.
In certain exceptional cases and with ores poor in sulphur, direct
blowing might be applicable, but for the general run of lead ores no
economy can be expected by doing away with the preliminary roast.
MAKING SULPHURIC ACID AT BROKEN HILL
(August 11, 1904)
The Broken Hill Proprietary Company has entered upon the
manufacture of sulphuric acid on a commercial scale. The acid is
practically a by-product, being made from the gases emanating
from the desulphurization of the ores, concentrates, etc., by the
Carmichael-Bradford process. The acid can be made at a minimum of
cost, and thus materially enhances the value of the process recently
introduced for the separation of zinc blende from the tailings by
flotation. The following particulars are taken from a recently
published description of the process: The ores, concentrates, slimes,
etc., as the case may be, are mixed with gypsum, the quantity of the
latter varying from 15 to 25 per cent. The mixture is then granulated
to the size of marbles and dumped into a converter. The bottom of
the charge is heated from 400 to 500 deg. C. It is then subjected to
an induced current of air, and the auxiliary heat is turned off. The
desulphurization proceeds very rapidly with the evolution of heat and
the gases containing sulphurous anhydride. The desulphurization is very
thorough, and no losses occur through volatilization. The sulphur thus
rendered available for acid making is rather more than is contained in
the ore, the sulphur in the agglomerated product being somewhat less
than that accounted for by the sulphur contained in the added gypsum.
Thus from one ton of 14 per cent. sulphide ore it is possible to make
about 12 cwt. of chamber acid, fully equaling 7 cwt. of strong acid.
The plant at present in use, which comprises a lead chamber of 40,000
cu. ft., can turn out 35 tons of chamber acid per week. This plant is
being duplicated, and it has also been decided to erect a large plant
at Port Pirie for use in the manufacture of superphosphates. It is
claimed that the production of sulphuric acid from ores containing only
14 per cent. of sulphur establishes a new record.
THE CARMICHAEL-BRADFORD PROCESS
BY DONALD CLARK
(November 3, 1904)
Subsequent to the introduction of the Huntington-Heberlein process
in Australia, Messrs. Carmichael and Bradford, two employees of the
Broken Hill Proprietary Company, patented a process which bears their
name. Instead of starting with lime, or limestone and galena, as in
the Huntington-Heberlein process, they discovered that if sulphate of
lime is mixed with galena and the temperature raised, on blowing a
current of air through the mixture the temperature rises and the mass
is desulphurized. The process would thus appear to be a corollary of
the original one, and the reactions in the converter are identical.
Owing to the success of the acid processes in separating zinc sulphide
from the tailing at Broken Hill, it became necessary to manufacture
sulphuric acid locally in large quantity. The Carmichael-Bradford
process has been started for the purpose of generating the sulphur
dioxide necessary, and is of much interest as showing how gases rich
enough in SO₂ may be produced from a mixture containing only from 13
to 16 per cent. sulphur.
Gypsum is obtained in a friable state within about five miles from
Broken Hill. This is dehydrated, the CaSO, 2H₂O being converted into
CaSO₄ on heating to about 200 deg. C. The powdered residue is mixed
with slime produced in the milling operations and concentrate in the
proportion of slime 3 parts, concentrate 1 part, and lime sulphate 1
part. The proportions may vary to some extent, but the sulphur contents
run from 13 to 16 or 17 per cent. The average composition of the
ingredients is as given in the table on the next page.
These materials are moistened with water and well mixed by passing
them through a pug-mill. The small amount of water used serves to
set the product, the lime sulphate partly becoming plaster of paris,
2CaSO, H₂O. While still moist the mixture is broken into pieces not
exceeding two inches in diameter and spread out on a drying floor,
where excess of moisture is evaporated by the conjoint action of sun
and wind.
─────────────────┬─────┬───────────┬────────┬────────
│SLIME│CONCENTRATE│CALCIUM │AVERAGE
│ │ │SULPHATE│
──────────────────┼─────┼───────────┼────────┼────────
Galena │ 24 │ 70 │ │ 29
Blende │ 30 │ 15 │ │ 21
Pyrite │ 3 │ │ │ 2
Ferric oxide │ 4 │ │ │ 2.5
Ferrous oxide │ 1 │ │ │ 1
Manganous oxide │ 6.5│ │ │ 5
Alumina │ 5.5│ │ │ 3
Lime │ 3.5│ │ 41 │ 10
Silica │ 23 │ │ │ 14
Sulphur trioxide │ │ │ 59 │ 12
─────────────────┴─────┴───────────┴────────┴────────
The pots used are small conical cast-iron ones, hung on trunnions,
and of the same pattern as used in the Huntington-Heberlein process.
Three of these are set in line, and two are at work while the third is
being filled. These pots have the same form of conical cover leading
to a telescopic tube, and all are connected to the same horizontal
pipe leading to the niter pots. Dampers are provided in each case. A
small amount of coal or fuel is fed into the pots and ignited by a
gentle blast; as soon as a temperature of about 400 to 500 deg. C. is
attained the dried mixture is fed in, until the pot is full; the cover
is closed down and the mass warms up. Water is first driven off, but
after a short time concentrated fumes of sulphur dioxide are evolved.
The amount of this gas may be as much as 14 per cent., but it is
usually kept at about 10 per cent., so as to have enough oxygen for
the conversion of the dioxide to the trioxide. The gases are led over
a couple of niter pots and thence to the usual type of lead chamber
having a capacity of 40,000 cu. ft. Chamber acid alone is made, since
this requires to be diluted for what is known as the saltcake process.
The plant has now been in operation for some time and, it is claimed,
with highly successful results. The product tipped out of. the
converter is similar to that obtained in the Huntington-Heberlein
process, and is at once fit for the smelters, the amount of sulphur
left in it being always less than that originally introduced with the
gypsum; analysis of the desulphurized material shows usually from 3 to
4 per cent. sulphur.
THE CARMICHAEL-BRADFORD PROCESS
BY WALTER RENTON INGALLS
(October 28, 1905)
As described in United States patent No. 705,904, issued July 29, 1902,
lead sulphide ore is mixed with 10 to 35 per cent. of calcium sulphate,
the percentage varying according to the grade of the ore. The mixture
is charged into a converter and gradually heated externally until the
lower portion of the charge, say one-third to one-fourth, is raised to
a dull-red heat; or the reactions may be started by throwing into the
empty converter a shovelful of glowing coal and turning on a blast of
air sufficient to keep the coal burning and then feeding the charge
on top of the coal. This heating effects a reaction whereby the lead
sulphide of the ore is oxidized to sulphate and the calcium sulphate is
reduced to sulphide. The heated mixture being continuously subjected
to the blast of air, the calcium sulphide is re-oxidized to sulphate
and is thus regenerated for further use. This reaction is exothermic,
and sufficient heat is developed to complete the desulphurization of
the charge of ore by the concurrent reactions between the lead sulphate
(produced by the calcium sulphate) and portions of undecomposed ore,
sulphurous anhydride being thus evolved. The various reactions, which
are complicated in their nature, continue until the temperature of
the charge reaches a maximum, by which time the charge has shrunk
considerably in volume and has a tendency to become pasty. This becomes
more marked as the production of lead oxide increases, and as the
desired point of desulphurization is attained the mixture fuses; at
this stage the calcium sulphide which is produced from the sulphate
cannot readily oxidize, owing to the difficulty of coming into actual
contact with the air in the pasty mass, but, being subjected to the
strong oxidizing effect of the metallic oxide, it is converted into
calcium plumbate, while sulphurous anhydride is set free. The mass then
cools, as the exothermic reactions cease, and can be readily removed to
a blast furnace for smelting.
The reactions above described are as outlined in the original
American patent specification. Irrespective of their accuracy,
the Carmichael-Bradford process is obviously quite similar to the
Huntington-Heberlein, and doubtless owes its origin to the latter. The
difference between them is that in the Huntington-Heberlein process
the ore is first partially roasted with addition of lime, and is then
converted in a special vessel. In the Carmichael-Bradford process
the ore is mixed with gypsum and is then converted directly. The
greatest claim for originality in the Carmichael-Bradford process
may be considered to lie in it as a method of desulphurizing gypsum,
inasmuch as not only is the sulphur of the ore expelled, but also a
part of the sulphur of the gypsum; and the sulphur is driven off as a
gas of sufficiently high tenor of sulphur dioxide to enable sulphuric
acid to be made from it economically. Up to the present time the
Carmichael-Bradford process has been put into practical use only at
Broken Hill, N. S. W.
The Broken Hill Proprietary Company first conducted a series of tests
in a converter capable of treating a charge of 20 cwt. These tests were
made at the smelting works at Port Pirie. Exhaustive experiments made
on various classes of ores satisfactorily proved the general efficacy
of the process. The following ores were tried in these preliminary
experiments, viz.:
First-grade concentrate containing: Pb, 60 per cent.; Zn, 10 per cent.;
S, 16 per cent.; Ag, 30 oz.
Second-grade concentrate containing: Pb, 45 per cent.; Zn, 12.5 per
cent.; S, 14.5 per cent.; Ag, 22 oz.
Slime containing: Pb, 21 per cent.; Zn, 17 per cent.; S, 13 per cent.;
Ag, 18 oz.
Lead-copper matte containing: Fe, 42 per cent.; Pb, 17 per cent.; Zn,
13.3 per cent.; Cu, 2.4 per cent.; S, 23 per cent.; Ag, 25 oz.
Other mattes, of varying composition up to 45 per cent. Pb and 100 oz.
Ag, were also tried.
The results from these preliminary tests were so gratifying that a
further set of tests was made on lead-zinc slime, with a view of
ascertaining whether any volatilization losses occurred during the
desulphurization. This particular material was chosen because of its
accumulation in large proportions at the mine, and the unsatisfactory
result of the heap roasting which has recently been practised. The
heap roasting, although affording a product containing only 7 per cent.
S, which is delivered in lump form and therefore quite suitable for
smelting, resulted in a high loss of metal by volatilization (17 per
cent. Pb, 5 per cent. Ag).
The result of nine charges of the slime treated by the
Carmichael-Bradford process was as follows:
─────────────────┬──────┬─────────────────────┬───────────────────────
│ │ ASSAYS │ CONTENTS
│ Cwt. ├────┬──────┬────┬────┼─────┬─────┬────┬──────
│ │Pb% │Ag oz.│Zn% │ S% │ Pb │ Ag. │ Zn │ S
│ │ │ │ │ │cwt. │ oz. │cwt.│cwt.
├──────┼────┼──────┼────┼────┼─────┼─────┼────┼──────
Raw slime │128.1 │21.3│ 18.0 │16.8│13.1│27.28│115.3│26.2│16.78
Raw gypsum │ 54.9 │ │ │ │ │ │ │ │ 9.88
├──────┤ │ │ │ ├─────┼─────┼────┼──────
Total │183.0 │ │ │ │ │27.28│115.3│25.2│26.66
──────────────────┼──────┼────┼──────┼────┼────┼─────┼─────┼────┼──────
Sintered material│109.88│20.7│ 17.2 │ │4.80│22.74│ 94.5│ │ 5.27
Middling │ 14.47│17.7│ 15.7 │ │6.20│ 2.56│ 11.3│ │ 0.89
Fines │ 11.12│19.0│ 14.8 │ │7.50│ 2.11│ 8.2│ │ 0.83
├──────┼────┼──────┼────┼────┼─────┼─────┼────┼──────
Total │135.47│ │ │ │5.17│27.41│113.0│ │ 6.99
─────────────────┴──────┴────┴──────┴────┴────┴─────┴─────┴────┴──────
These results indicated practically no volatilization of lead and
silver during the treatment, the lead showing a slight increase, viz.,
0.47 per cent., and the silver 1.13 per cent. loss. A desulphurization
of 70.4 per cent. was effected. A higher desulphurization could have
been effected had this been desired. In the above tabulated results,
the term “middling” is applied to the loose fritted lumps lying on the
top of the charge: these are suitable for smelting, the fines being the
only portion which has to be returned.
In order to test the practicability of making sulphuric acid, a plant
consisting of three large converters of capacity of five tons each,
together with a lead chamber 100 ft. by 20 ft. by 20 ft., was then
erected at Broken Hill, together with a dehydrating furnace, pug-mill,
and granulator. These converters are shown in the accompanying
engravings.
A trial run was made with 108 tons of concentrate of the following
composition: 54 per cent. lead; 1.9 per cent. iron; 0.9 per cent.
manganese; 9.4 per cent. zinc; 14.6 per cent. sulphur; 19.2 per cent.
insoluble residue, and 24 oz. silver per ton.
The converter charge consisted of 100 parts of the concentrate and
25 parts of raw gypsum, crushed to pass a 1 in. hole and retained
by a 0.25 in. hole, the material finer than 0.25 in. (which amounted
to 5 per cent. of the total) being returned to the pug-mill. After
desulphurization in the converter, the product assayed as follows:
48.9 per cent. lead; 1.80 per cent. iron; 0.80 per cent. manganese;
7.87 per cent. zinc; 3.90 per cent. sulphur; 1.02 per cent. alumina;
5.80 per cent. lime; 21.75 per cent. insoluble residue; 8.16 per cent.
undetermined (oxygen as oxides, sulphates, etc.); total, 100 per cent.
Its silver content was 22 oz. per ton. The desulphurized ore weighed
10 per cent. more than the raw concentrate. During this run 34 tons of
acid were made.
A trial was then made on 75 tons of slime of the following composition:
18.0 per cent. lead; 16.6 per cent. zinc; 6.0 per cent. iron; 2.5 per
cent. manganese; 3.2 per cent. alumina; 2.1 per cent. lime; 38.5 per
cent. insoluble residue; total, 100 per cent. Its silver content was
19.2 oz. per ton.
The converter charge in this case consisted of 100 parts of raw slime
and 30 parts of gypsum. The converted material assayed as follows:
16.1 per cent. lead; 14.0 per cent. zinc; 3.6 per cent. sulphur; 5.42
per cent. iron; 2.25 per cent. manganese; 4.10 per cent. alumina; 8.60
per cent. lime; 39.80 per cent. insoluble residue; 6.13 per cent.
undetermined (oxygen, etc.); total, 100 per cent.; and silver, 17.5
oz. per ton. The increase in weight of desulphurized ore over that
of the raw ore was 11 per cent. During this run 22 tons of acid were
manufactured.
The analysis of the gypsum used in each of the above tests (at Broken
Hill) was as follows: 76.1 per cent. CaSO₄, 2H₂O; 0.5 per cent.
Fe₂O₃; 4.5 per cent. Al₂O₃; 18.9 per cent. insoluble
residue.
The plant was then put into continuous operation on a mixture of three
parts slime and one of concentrate, desulphurizing down to 4 per cent.
S, and supplying 20 tons of acid per week, and additions were made to
the plant as soon as possible. The acid made at Broken Hill has been
used in connection with the Delprat process for the concentration of
the zinc tailing. At Port Pirie, works are being erected with capacity
for desulphurization of about 35,000 tons per annum, with an acid
output of 10,000 tons. This acid is to be utilized for the acidulation
of phosphate rock.
[Illustration: FIG. 15.—Details of Converters.]
The cost of desulphurization of a ton of galena concentrate by the
Carmichael-Bradford process, based on labor at $1.80 per 8 hours,
gypsum at $2.40 per 2240 lb., and coal at $8.40 per 2240 lb., is
estimated as follows:
0.25 ton of gypsum $0.60
Dehydrating and granulating gypsum .48
Drying mixture of ore and gypsum .12
Converting .24
Spalling sintered material .12
0.01 ton coal .08
——-——-
Total $1.64
The lime in the sintered product is credited at 12c., making the net
cost $1.52 per ton (2240 lb.) of ore.
The plant required for the Carmichael-Bradford process can be described
with sufficient clearness without drawings, except the converters. The
ore (concentrate, slime, etc.) to be desulphurized is delivered at the
top of the mill by cars, conveyors, or other convenient means, and
dumped into a bin. Two screw feeders placed inside the bin supply the
mill with ore, uniformly and as fast as it is required. These feeders
deliver the ore into a chute, which directs it into a vertical dry
mixer.
A small bin, on the same level as the ore-bin, receives the crude
gypsum from cars. Thence it is fed automatically to a disintegrator,
which pulverizes it finely and delivers it into a storage bin
underneath. This disintegrator revolves at about 1700 r.p.m. and
requires 10 h.p. The body of the machine is cast iron, fitted with
renewable wearing plates (made of hard iron) in the grinding chamber.
The revolving parts consist of a malleable iron disc in which are fixed
steel beaters, faced on the grinding surface with highly tempered
steel. The bin that receives the floured gypsum contains a screw
conveyor similar to those in the ore-bin, and dumps the material into
push conveyors passing into the dehydrating furnace. They carry the
crushed gypsum along at a speed of about 1 ft. per minute, and allow
about 20 ft. to dehydrate the gypsum. This speed can, of course, be
regulated to suit requirements.
The dehydrated gypsum runs down a chute into an elevator boot, and is
elevated into a bin which is on the same level as the ore-bin. This bin
also contains a screw conveyor, like that in the ore-bin. The speed of
delivery is regulated to deliver the right proportion of dehydrated
gypsum to the mixer.
The mixer is of the vertical pattern and receives the sulphide ore
and dehydrated gypsum from the screw feeders. In it are set two flat
revolving cones running at different speeds, thus ensuring a thorough
mixture of the gypsum and ore. The mixed material drops from the
cones upon two baffle plates, and is wetted just before entering the
pug-mill. The pug-mill is a wrought-iron cylinder of ¼ in. plate about
2 ft. 6 in. diameter and 6 or 8 ft. long, and has the mixer fitted
to the head. It contains a 3 ft. wrought-iron spiral with propelling
blades, which forces the plastic mixture through ¾ in. holes in the
cover. The material comes out in long cylindrical pieces, but is broken
up and formed into marble-shaped pieces on dropping into a revolving
trommel.
The trommel is about 5 ft. long, 2 ft. in diameter at the small end and
about 4 ft. at the large end. It revolves about a wrought-iron spindle
(2½ in. diameter) carrying two cast-iron hubs to which are fitted
arms for carrying the conical plate ⅛ in. thick. About 18 in. of
the small end of the cone is fitted with wire gauze, so as to prevent
the material as it comes out of the pug-mill from sticking to it. The
trommel is driven by bevel gearing at 20 to 25 r.p.m. The granulated
material formed in the trommel is delivered upon a drying conveyor.
The conveyor consists of hinged wrought-iron plates flanged at the side
to keep the material from running off. It is driven from the head by
gearing, at a speed of 1 ft. per minute, passing through a furnace 10
ft. long to dry and set the granules of ore and gypsum. This speed can,
of course, be regulated to suit requirements. The granulated material,
after leaving the furnace, is delivered to a single-chain elevator,
traveling at a speed of about 150 ft. per minute. It drops the material
into a grasshopper conveyor, driven by an eccentric, which distributes
the material over the length of a storage bin. From this bin the
material is directed into the converters by means of the chutes, which
have their bottom ends hinged so as to allow for the raising of the
hood when charging the converters.
The converters are shown in the accompanying engravings, but they may
be of slightly different form from what is shown therein, i.e., they
may be more spherical than conical. They will have a capacity of about
four tons, being 6 ft. in diameter at the top, 4 ft. in diameter at
the false bottom, and about 5 ft. deep. They are swung on cast-iron
trunnions bolted to the body and turned by means of a hand-wheel and
worm (not shown). They are carried on strong cast-iron standards fitted
with bearings for trunnions, and all necessary brackets for tilting
gear. The hood has a telescopic funnel which allows it to be raised
or lowered, weights being used to balance it. At the apex of the cone
a damper is provided to regulate the draft. A 4 in. hole in the pot
allows the air from the blast-pipe, 18 in. in diameter, to enter under
the false perforated bottom, the connection between the two being made
by a flexible pipe and coupling. Two Baker blowers supply the blast for
the converters. The material, after being sintered, is tipped on the
floor in front of the converters and is there broken up to any suitable
size, and thence dispatched to the smelters.
[Illustration: FIG. 16.—Arrangement of Converters.]
The necessary power for a plant with a capacity of 150 tons of ore per
day will be supplied by a 50 h.p. engine.
THE SAVELSBERG PROCESS
BY WALTER RENTON INGALLS
(December 9, 1905)
There are in use at the present time three processes for the
desulphurization of galena by the new method, which has been referred
to as the “lime-roasting of galena.” The Huntington-Heberlein and the
Carmichael-Bradford processes have been previously described. The third
process of this type, which in some respects is more remarkable than
either of the others, is the invention of Adolf Savelsberg, director
of the smeltery at Ramsbeck, Westphalia, Germany, which is owned by
the Akt. Gesell. f. Bergbau, Blei. u. Zinkhüttenbetrieb zu Stolberg
u. in Westphalen. The process is in use at the Ramsbeck and Stolberg
lead smelteries of that company. It is described in American patent
No. 755,598, issued March 22, 1904 (application filed Dec. 18, 1903).
The process is well outlined in the words of the inventor in the
specification of that patent:
“The desulphurizing of certain ores has been effected by blowing air
through the ore in a chamber, with the object of doing away with the
imperfect and costly process of roasting in ordinary furnaces; but
hitherto it has not been possible satisfactorily to desulphurize lead
ores in this manner, as, if air be blown through raw lead ores in
accordance with either of the processes used for treating copper ores,
for example, the temperature rises so rapidly that the unroasted lead
ore melts and the air can no longer act properly upon it, because
by reason of this melting the surface of the ores is considerably
decreased, the greater number of points or extent of surface which
the raw ore originally presented to the action of the oxygen of the
air blown through being lost, and, moreover, the further blowing
of air through the molten mass of ore produces metallic lead and a
plumbiferous slag (in which the lead oxide combines with the gangue)
and also a large amount of light dust, consisting mainly of sublimated
lead sulphide. Huntington and Heberlein have proposed to overcome
these objections by adopting a middle course, consisting in roasting
the ores with the addition of limestone for overcoming the ready
fusibility of the ores, and then subjecting them to the action of the
current of air in the chamber; but this process is not satisfactory,
because it still requires the costly previous operation in a roasting
furnace.
[Illustration: Fig. 18.—Converter Ready to Dump.]
“My invention is based on the observation which I have made that if
the lead ores to be desulphurized contain a sufficient quantity of
limestone it is possible, by observing certain precautions, to dispense
entirely with the previous roasting in a roasting furnace, and to
desulphurize the ores in one operation by blowing air through them. I
have found that the addition of limestone renders the roasting of the
lead ore unnecessary, because the limestone produces the following
effects:
“The particles of limestone act mechanically by separating the
particles of lead ore from each other in such a way that premature
agglomeration is prevented and the whole mass is loosened and rendered
accessible to air; and, moreover, the limestone moderates the high
reaction temperature resulting from the burning of the sulphur, so
that the liquefaction of the galena, the sublimation of lead sulphide,
and the separation of metallic lead are avoided. The moderation of
the temperature of reaction is caused by the decomposition of the
limestone into caustic lime and carbon dioxide, whereby a large amount
of heat becomes latent. Further, the decomposition of the limestone
causes chemical reactions, lime being formed, which at the moment of
its formation is partly converted into sulphate of lime at the expense
of the sulphur contained in the ore, and this sulphate of lime, when
the scorification takes place, is transformed into calcium silicate
by the silicic acid, the sulphuric acid produced thereby escaping.
The limestone also largely contributes to the desulphurization of the
ore, as it causes the production of sulphuric acid at the expense of
the sulphur of the ore, which sulphuric acid is a powerful oxidizing
agent. If, therefore, a mixture of raw lead ore and limestone (which
mixture must, of course, contain a sufficient amount of silicic acid
for forming silicates) be introduced into a chamber and a current of
air be blown through the mixture, and at the same time the part of the
mixture which is near the blast inlet be ignited, the combustion of the
sulphur will give rise to very energetic reactions, and sulphurous
acid, sulphuric acid, lead oxide, sulphates and silicates are produced.
The sulphurous acid and the carbon dioxide escape, while the sulphuric
acid and sulphates act in their turn as oxidizing agents on the
undecomposed galena. Part of the sulphates is decomposed by the silicic
acid, thereby liberating sulphuric acid, which, as already stated, acts
as an oxidizing agent. The remaining lead oxide combines finally with
the gangue of the ore and the non-volatile constituents of the flux
(the limestone) to form the required slag. These several reactions
commence at the blast inlet at the bottom of the chamber, and extend
gradually toward the upper portion of the charge of ore and limestone.
Liquefaction of the ores does not take place, for although a slag is
formed it is at once solidified by the blowing in of the air, the
passages formed thereby in the hardening slag allowing of the continued
passage therethrough of the air. The final product is a silicate
consisting of lead oxide, lime, silicic acid, and other constituents of
the ore, which now contains but little or no sulphur and constitutes a
coherent solid mass, which, when broken into pieces, forms a material
suitable to be smelted.
“The quantity of limestone required for the treatment of the lead
ores varies according to the constitution of the ores. It should,
however, amount generally to from 15 to 20 per cent. As lead ores do
not contain the necessary amount of limestone as a natural constituent,
a considerable amount of limestone must be added to them, and this
addition may be made either during the dressing of the ores or
subsequently.
“For the satisfactory working of the process, the following precautions
are to be observed: In order that the blowing in of the air may not
cause particles of limestone to escape in the form of dust before
the reaction begins, it is necessary to add to the charge before it
is subjected to the action in the chamber a considerable amount of
water—say 5 per cent. or more. This water prevents the escape of dust,
and it also contributes considerably to the formation of sulphuric
acid, which, by its oxidizing action, promotes the reaction, and,
consequently, also the desulphurization. It is advisable, in conducting
the operation, not to fill the chamber with the charge at once, but
first only partly to fill it and add to the charge gradually while the
chamber is at work, as by this means the reaction will take place more
smoothly in the mass.
[Illustration: Fig. 19.—Charge Dumped.]
“It is advantageous to proceed as follows: The bottom part of a
chamber of any suitable form is provided with a grate, on which is
laid and ignited a mixture of fuel (coal, coke, or the like) and
pieces of limestone. By mixing the fuel with pieces of limestone the
heating power of the fuel is reduced and the grate is protected,
while at the same time premature melting of the lower part of the
charge is prevented; or the grate may be first covered with a layer
of limestone and the fuel be laid thereon, and then another layer of
limestone be placed on the fuel. On the material thus placed in the
chamber, a uniform charge of lead ore and limestone—say about 12 in.
high—is placed, this having been moistened as previously explained.
Under the influence of the air-blast and the heat, the reactions
hereinbefore described take place. When the upper surface of the first
layer becomes red-hot, a further charge is laid thereon, and further
charges are gradually introduced as the surface of the preceding
charge becomes red-hot, until the chamber is full. So long as charges
are still introduced a blast of air of but low pressure is blown
through; but when the chamber is filled a larger quantity of air at a
higher pressure is blown through. The scorification process then takes
place, a very powerful desulphurization having preceded it. During the
scorification the desulphurization is completed.
“When the process is completed, the chamber is tilted and the
desulphurized mass falls out and is broken into small pieces for
smelting.”
The drawing on page 190, Fig. 17, shows a side view of the apparatus
used in connection with the process, which will be readily understood
without special description. The dotted lines show the pot in its
emptying position. The series of operations is clearly illustrated in
Figs. 18-20, which are reproduced from photographs.
This process has now been in practical use at Ramsbeck for three years,
where it is employed for the desulphurization of galena of high grade
in lead, with which are mixed quartzose silver ore (or sand if no such
ore be available), and calcareous and ferruginous fluxes. A typical
charge is 100 parts of lead ore, 10 parts of quartzose silver ore,
10 parts of spathic iron ore, and 19 parts of limestone. A thorough
mixture of the components is essential; after the mixture has been
effected, the charge is thoroughly wetted with about 5 per cent.
of water, which is conceived to play a threefold function in the
desulphurizing operation, namely: (1) preservation of the homogeneity
of the mixture during the blowing; (2) reduction of temperature during
the process; and (3) formation of sulphuric acid in the process, which
promotes the desulphurization of the ore.
[Illustration: FIG. 17.—Savelsberg Converter.]
The moistened charge is conveyed to the converters, into which it
is fed in thin layers. The converters are hemispherical cast-iron
pots, supported by trunnions on a truck, as shown in the accompanying
engravings. Except for this method of support, which renders the
pot movable, the arrangement is quite similar to that which is
employed in the Huntington-Heberlein process. The pots which are now
in use at Ramsbeck have capacity for about 8000 kg. of charge, but
it is the intention of the management to increase the capacity to
10,000 or 12,000 kg. Previously, pots of only 5000 kg. capacity were
employed. Such a pot weighed 1300 kg., exclusive of the truck. The
air-blast was about 7 cu. m. (247.2 cu. ft.) per min., beginning at
a pressure of 10 to 20 cm. of water (2¾ to 4½ oz.) and rising to
50 to 60 cm. (11½ to 13½ oz.) when the pot was completely filled with
charge. The desulphurization of a charge is completed in 18 hours. A
pot is attended by one man per shift of 12 hours; this is only the
attention of the pot proper, the labor of conveying material to it and
breaking up the desulphurized product being extra. One man per shift
should be able to attend to two pots, which is the practice in the
Huntington-Heberlein plants.
[Illustration: Fig. 20.—Converter in Position for Blowing.]
When the operation in the pot is completed, the latter is turned on its
trunnions, until the charge slides out by gravity, which it does as a
solid cake. This is caused to fall upon a vertical bar, which breaks
it into large pieces. By wedging and sledging these are reduced to
lumps of suitable size for the blast furnace. When the operation has
been properly conducted the charge is reduced to about 2 or 3 per cent.
sulphur. It is expected that the use of larger converters will show
even more favorable results in this particular.
As in the Huntington-Heberlein and Carmichael-Bradford processes, one
of the greatest advantages of the Savelsberg process is the ability to
effect a technically high degree of desulphurization with only a slight
loss of lead and silver, which is of course due to the perfect control
of the temperature in the process. The precise loss of lead has not yet
been determined, but in the desulphurization of galena containing 60
to 78 per cent. lead, the loss of lead is probably not more than 1 per
cent. There appears to be no loss of silver.
The process is applicable to a wide variety of lead-sulphide ores. The
ore treated at Ramsbeck contains 60 to 78 per cent. lead and about
15 per cent. of sulphur, but ore from Broken Hill, New South Wales,
containing 10 per cent. of zinc has also been treated. A zinc content
up to 7 or 8 per cent. in the ore is no drawback, but ores carrying a
higher percentage of zinc require a larger addition of silica and about
5 per cent. of iron ore in order to increase the fusibility of the
charge. The charge ordinarily treated at Ramsbeck is made to contain
about 11 per cent. of silica. The presence of pyrites in the ore is
favorable to the desulphurization. Dolomite plays the same part in
the process that limestone does, but is of course less desirable, in
view of the subsequent smelting in the blast furnace. The ore is best
crushed to about 3 mm. size, but good results have been obtained with
ore coarser in size than that. However, the proper size is somewhat
dependent upon the character of the ore. The blast pressure required in
the converter is also, of course, somewhat dependent upon the porosity
of the charge. Fine slimes are worked up by mixture with coarser ore.
In making up the charge, the proportion of limestone is not varied
much, but the proportions of silica and iron must be carefully modified
to suit the ore. Certain kinds of ore have a tendency to remain
pulverulent, or to retain balls of unsintered, powdered material.
In such cases it is necessary to provide more fusible material in
the charge, which is done by varying the proportions of silica and
iron. The charge must, moreover, be prepared in such a manner that
overheating, and consequently the troublesome fusion of raw galena,
will be avoided.
The essential difference between the Huntington-Heberlein and
Savelsberg processes is the use in the former of a partially
desulphurized ore, containing lime and sulphate of lime; and the use
in the latter of raw ore and carbonate of lime. It is claimed that the
latter, which loses its carbon dioxide in the converter, necessarily
plays a different chemical part from that of quicklime or gypsum.
Irrespective of the reactions, however, the Savelsberg process has the
great economic advantage of dispensing with the preliminary roasting of
the Huntington-Heberlein process, wherefore it is cheaper both in first
cost of plant and in operation.
THE LIME-ROASTING OF GALENA[32]
BY WALTER RENTON INGALLS
During the last two years, and especially during the last six
months, a number of important articles upon the new methods for the
desulphurization of galena have been published in the technical
periodicals, particularly in the _Engineering and Mining Journal_
and in _Metallurgie_. I proposed for these methods the type-name
of “lime-roasting of galena,” as a convenient metallurgical
classification,[33] and this term has found some acceptance. The
articles referred to have shown the great practical importance of these
new processes, and the general recognition of their metallurgical and
commercial value, which has already been accorded to them. It is my
present purpose to review broadly the changes developed by them in
the metallurgy of lead, in which connection it is necessary to refer
briefly to the previous state of the art.
The elimination of the sulphur content of galena has been always the
most troublesome part of the smelting process, being both costly in the
operation and wasteful of silver and lead. Previous to the introduction
of the Huntington-Heberlein process at Pertusola, Italy, it was
effected by a variety of methods. In the treatment of non-argentiferous
galena concentrate, the smelting was done by the roast-reduction method
(roasting in reverberatory furnace and smelting in blast furnace);
the roast-reaction method, applied in reverberatory furnaces; and the
roast-reaction method, applied in Scotch hearths.[34] Precipitation
smelting, simple, had practically gone out of use, although its
reactions enter into the modern blast-furnace practice, as do also
those of the roast-reaction method.
In the treatment of argentiferous lead ores, a combination of the
roast-reduction, roast-reaction and precipitation methods had been
developed. Ores low in lead were still roasted, chiefly in hand-worked
reverberatories (the mechanical furnaces not having proved well adapted
to lead-bearing ores), while the high loss of lead and silver in
sinter-or slag-roasting of rich galenas had caused those processes to
be abandoned, and such ores were charged raw into the blast furnace,
the part of their sulphur which escaped oxidation therein reappearing
in the form of matte. In the roast-reduction smelting of galena alone,
however, there was no way of avoiding the roasting of the whole, or at
least a very large percentage of the ore, and in this roasting the ore
had necessarily to be slagged or sintered in order to eliminate the
sulphur to a satisfactory extent. This is exemplified in the treatment
of the galena concentrate of southeastern Missouri at the present time.
Until the two new Scotch-hearth plants at Alton and Collinsville, Ill.,
were put in operation, the three processes of smelting the southeastern
Missouri galena were about on an equal footing. Their results per ton
of ore containing 65 per cent. lead were approximately as follows[35]:
──────────────────┬──────────────┬────────────
METHOD │ COST │ EXTRACTION
──────────────────┼──────────────┼───────────
Reverberatory │ $6.50-7.00 │ 90-92%
Scotch hearth │ 5.75-6.50 │ 87-88%
Roast-reduction │ 6.00-7.00 │ 90-92%
──────────────────┴──────────────┴───────────
The new works employ the Scotch-hearth process, with bag-houses for
the recovery of the fume, which previously was the weak point of this
method of smelting.[36] This improvement led to a large increase in the
recovery of lead, so that the entire extraction is now approximately 98
per cent. of the content of the ore, while on the other hand the cost
of smelting per ton of ore has been reduced through the increased size
of these plants and the introduction of improved means for handling
ore and material. The practice of these works represents the highest
efficiency yet obtained in this country in the smelting of high-grade
galena concentrate, and probably it cannot be equaled even by the
Huntington-Heberlein and similar processes. The Scotch-hearth and
bag-house process is therefore the one of the older methods of smelting
which will survive.
In the other methods of smelting, a large proportion of the cost is
involved in the roasting of the ore, which amounts in hand-worked
reverberatory furnaces to $2 to $2.50 per ton. Also, the larger
proportion of the loss of metal is suffered in the roasting of the ore,
this loss amounting to from 6 to 8 per cent. of the metal content of
such ore as is roasted. The loss of lead in the combined process of
treatment depends upon the details of the process. The chief advantage
of lime-roasting in the treatment of this class of ore is in the higher
extraction of metal which it affords. This should rise to 98 per cent.
That figure has been, indeed, surpassed in operations on a large scale,
extending over a considerable period.
In the treatment of the argentiferous ores of the West different
conditions enter into the consideration. In the working of those ores,
the present practice is to roast only those which are low in lead,
and charge raw into the blast furnace the rich galenas. The cost of
roasting is about $2 to $2.50 per ton; the cost of smelting is about
$2.50 per ton. On the average about 0.4 ton of ore has to be roasted
for every ton that is smelted. The cost of roasting and smelting is
therefore about $3.50 per ton. In good practice the recovery of silver
is about 98 per cent. and of lead about 95 per cent., reckoned on basis
of fire assays.
In treatment of these ores, the lime-roasting process offers several
advantages. It may be performed at less than the cost of ordinary
roasting.[37] The loss of silver and lead during the roasting is
reduced to insignificant proportion. The sulphide fines which must be
charged raw into the blast furnace are eliminated, inasmuch as they
can be efficiently desulphurized in the lime-roasting pots without
significant loss; all the ore to be smelted in the blast furnace
can be, therefore, delivered to it in lump form, whereby the speed
of the blast furnace is increased and the wind pressure required
is decreased. Finally, the percentage of sulphur in the charge is
reduced, producing a lower matte-fall, or no matte-fall whatever, with
consequent saving in expense of retreatment. In the case of a new
plant, the first cost of construction and the ground-space occupied
are materially reduced. Before discussing more fully the extent and
nature of these savings, it is advisable to point out the differences
among the three processes of lime-roasting that have already come into
practical use.
In the Huntington-Heberlein process, the ore is mixed with suitable
proportions of limestone and silica (or quartzose ore) and is then
partially roasted, say to reduction of the sulphur to one half. The
roasting is done at a comparatively low temperature, and the loss of
metals is consequently small. The roasted ore is dampened and allowed
to cool. It is then charged into a hemispherical cast-iron pot, with
a movable hood which covers the top and conveys off the gases. There
is a perforated grate in the bottom of the pot, on which the ore
rests, and air is introduced through a pipe entering the bottom of the
pot, under the grate. A small quantity of red-hot calcines from the
roasting furnaces is thrown on the grate to start the reaction; a layer
of cold, semi-roasted ore is put upon it, the air blast is turned on
and reaction begins, which manifests itself by the copious evolution
of sulphur fumes. These consist chiefly of sulphur dioxide, but they
contain more or less trioxide, which is evident from the solution of
copperas that trickles from the hoods and iron smoke-pipes, wherein the
moisture condenses. As the reaction progresses, and the heat creeps
up, more ore is introduced, layer by layer, until the pot is full.
Care is taken by the operator to compel the air to pass evenly and
gently through the charge, wherefore he is watchful to close blow-holes
which develop in it. At the end of the operation, which may last from
four to eighteen hours, the ore becomes red-hot at the top. The hood
is then pushed up, and the pot is turned on its trunnions, by means
of a hand-operated wheel and worm-gear, until the charge slides out,
which it does as a solid, semi-fused cake. The pot is then turned back
into position. Its design is such that the air-pipe makes automatic
connection, a flanged pipe cast with the pot settling upon a similarly
flanged pipe communicating with the main, a suitable gasket serving
to make a tight joint. The pots are set at an elevation of about 12
ft. above the ground, so that when the charge slides out the drop will
break it up to some extent, and it is moreover caused to fall on a
wedge, or similar contrivance, to assist the breakage. After cooling it
is further broken up to furnace size by wedging and sledging; the lumps
are forked out, and the fines screened and returned to a subsequent
charge for completion of their desulphurization.
The Savelsberg process differs from the Huntington-Heberlein in respect
to the preliminary roasting, which in the Savelsberg process is
omitted, the raw ore, mixed with limestone and silica, being charged
directly into the converter. The Savelsberg converter is supported on
a truck, instead of being fixed in position, but otherwise its design
and management are quite similar to those of the Huntington-Heberlein
converter. In neither case are there any patents on the converters.
The patents are on the processes. In view of the litigation that
has already been commenced between their respective owners, it is
interesting to examine the claims.
The Huntington-Heberlein patent (U. S. 600,347, issued March 8, 1898,
applied for Dec. 9, 1896) has the following claims:
1. The herein-described method of oxidizing sulphide ores of lead
preparatory to reduction to metal, which consists in mixing with the
ore to be treated an oxide of an alkaline-earth metal, such as calcium
oxide, subjecting the mixture to heat in the presence of air, then
reducing the temperature and finally passing air through the mass
to complete the oxidation of the lead, substantially as and for the
purpose set forth.
2. The herein-described method of oxidizing sulphide ores of lead
preparatory to reduction to metal, which consists in mixing calcium
oxide or other oxide of an alkaline-earth metal with the ore to be
treated, subjecting the mixture in the presence of air to a bright-red
heat (about 700 deg. C.), then cooling down the mixture to a dull-red
heat (about 500 deg. C.), and finally forcing air through the mass
until the lead ore, reduced to an oxide, fuses, substantially as set
forth.
3. The herein-described method of oxidizing lead sulphide in the
preparation of the same for reduction to metal, which consists in
subjecting the sulphide to a high temperature in the presence of an
oxide of an alkaline-earth metal, such as calcium oxide, and oxygen,
and then lowering the temperature substantially as set forth.
Adolf Savelsberg, in U. S. patent 755,598 (issued March 22, 1904,
applied for Dec. 18, 1903) claims:
1. The herein-described process of desulphurizing lead ores, which
consists in mixing raw ore with limestone and then subjecting the
mixture to the simultaneous application of heat and a current of air in
sufficient proportions to substantially complete the desulphurization
in one operation, substantially as described.
2. The herein-described process of desulphurizing lead ores, which
process consists in first mixing the ores with limestone, then
moistening the mixture, then filling it without previous roasting into
a chamber, then heating it and treating it by a current of air, as and
for the purpose described.
3. The herein-described process of desulphurizing lead ores, which
consists in mixing raw ores with limestone, then filling the mixture
into a chamber, then subjecting the mixture to the simultaneous
application of heat and a current of air in sufficient proportions
to substantially complete the desulphurization in one operation, the
mixture being introduced into the chamber in partial charges introduced
successively at intervals during the process, substantially as
described.
4. The herein-described process of desulphurizing lead ores, then
moistening the mixture, then filling it without previous roasting into
a chamber, then heating it and treating it by a current of air, the
mixture being introduced into the chamber in partial charges introduced
successively at intervals during the process, as and for the purpose
described.
5. The herein-described process of desulphurizing lead ores, which
process consists in first mixing the ores with sufficient limestone to
keep the temperature of the mixture below the melting-point of the ore,
then filling the mixture into a chamber, then heating said mixture and
treating it with a current of air, as and for the purpose described.
6. The herein-described process of desulphurizing lead ores, which
process consists in first mixing the ores with sufficient limestone to
mechanically separate the particles of galena sufficiently to prevent
fusion, and to keep the temperature below the melting-point of the ore
by the liberation of carbon dioxide, then filling the mixture into a
chamber, then heating said mixture and treating it with a current of
air, as and for the purpose described.
The Carmichael-Bradford process differs from the Savelsberg by the
treatment of the raw ore mixed with gypsum instead of limestone,
and differs from the Huntington-Heberlein both in respect to the
use of gypsum and the omission of the preliminary roasting. The
Carmichael-Bradford process has not been threatened with litigation,
so far as I am aware. The claims of its original patent read as
follows[38]:
1. The process of treating mixed sulphide ores, which consists in
mixing with said ores a sulphur compound of a metal of the alkaline
earths, starting the reaction by heating the same, thereby oxidizing
the sulphide and reducing the sulphur compound of the alkali metal,
passing a current of air to oxidize the reduced sulphide compound of
the metal of the alkalies preparatory to acting upon a new charge of
sulphide ores, substantially as and for the purpose set forth.
2. The process of treating mixed sulphide ores, which consists in
mixing calcium sulphate with said ores, starting the reaction by
means of heat, thereby oxidizing the sulphide ores, liberating
sulphurous-acid gas and converting the calcium sulphate into calcium
sulphide and oxidizing the calcium sulphide to sulphate preparatory to
treating a fresh charge of sulphide ores, substantially as and for the
purpose set forth.
The process described by W. S. Bayston, of Melbourne (Australian patent
No. 2862), appears to be identical with that of Savelsberg.
Irrespective of the validity of the Savelsberg and Carmichael-Bradford
patents, and without attempting to minimize the ingenuity of their
inventors and the importance of their discoveries, it must be conceded
that the merit for the invention and introduction of lime-roasting of
galena belongs to Thomas Huntington and Ferdinand Heberlein. The former
is an American, and this is the only claim that the United States can
make to a share in this great improvement in the metallurgy of lead. It
is to be regretted, moreover, that of all the important lead-smelting
countries in the world, America has been the most backward in adopting
it.
The details of the three processes and the general results accomplished
by them have been rather fully described in a series of articles
recently published in the _Engineering and Mining Journal_. There
has been, however, comparatively little discussion as to costs; and
unfortunately the data available for analysis are extremely scanty, due
to the secrecy with which the Huntington-Heberlein process, the most
extensively exploited of the three, has been veiled. Nevertheless, I
may attempt an approximate estimation of the various details, taking
the Huntington-Heberlein process as the basis.
The ore, limestone and silica are crushed to pass a four-mesh screen.
This is about the size to which it would be necessary to crush as
preliminary to roasting in the ordinary way, wherefore the only
difference in cost is the charge for crushing the limestone and silica,
which in the aggregate may amount to one-sixth of the weight of the raw
sulphide and may consequently add 2 to 2.5c. to the cost of treating
a ton of ore. The mixing of ore and fluxes may be costly or cheap,
according to the way of doing it. If done in a rational way it ought
not to cost more than 10c. per ton of ore, and may come to less. The
delivery of the ore from the mixing-house to the roasting furnaces
ought to be done entirely by mechanical means, at insignificant cost.
The Heberlein roasting furnace, which is used in connection with the
H.-H. process, is simply an improvement on the old Brunton calciner—a
circular furnace, with revolving hearth. The construction of this
furnace, according to American designs, is excellent. The hearth is
26 ft. in diameter; it is revolved at slow speed and requires about
1.5 h.p. A flange at the periphery of the hearth dips into sand in an
annular trough, thus shutting off air from the combustion chamber,
except through the ports designed for its admittance. The mechanical
construction of the furnace is workmanlike, and the mechanism under the
hearth is easy of access and comfortably attended to.
A 26 ft. furnace roasts about 80,000 lb. of charge per 24 hours. In
dealing with an ore containing 20 to 22 per cent. of sulphur, the
latter is reduced to about 10 to 11 per cent., the consumption of
coal being about 22.5 per cent. of the weight of the charge. The
hearth efficiency is about 150 lb. per sq. ft., which in comparison
with ordinary roasting is high. The coal consumption, however, is not
correspondingly low. Two furnaces can be managed by one man per 8 hour
shift. On the basis of 80 tons of charge ore per 24 hours, the cost of
roasting should be approximately as follows:
Labor—3 men at $2.50 $ 7.50
Coal—18 tons at $2 36.00
Power 3.35
Repairs 3.35
——————
Total $50.20 = 63c. per ton.
In the above estimate repairs have been reckoned at the same figure as
is experienced with Brückner cylinders, and the cost of power has been
allowed for with fair liberality. The estimated cost of 63c. per ton
is comparable with the $1.10 to $1.45 per ton, which is the result of
roasting in Brückner cylinders in Colorado, reducing the ore to 4.5-6
per cent. sulphur.
The Heberlein furnace is built up to considerable elevation above
the ground level, externally somewhat resembling the Pearce turret
furnace. This serves two purposes: (1) it affords ample room under the
hearth for attention to the driving mechanism; and (2) it enables the
ore to be discharged by gravity into suitable hoppers, without the
construction of subterranean gangways. The ore discharges continuously
from the furnace, at dull-red heat, into a brick bin, wherein it is
cooled by a water-spray. Periodically a little ore is diverted into a
side bin, in which it is kept hot for starting a subsequent charge in
the converter.
The cooled ore is conveyed from the receiving bins at the roasting
furnaces to hopper-bins above the converters. If the tramming be done
by hand the cost, with labor at 25c. per hour, may be approximately
12.5c. per ton of ore, but this should be capable of considerable
reduction by mechanical conveyance.
The converters are hemispherical pots of cast iron, 9 ft. in
diameter at the top, and about 4 ft. in depth. They are provided
with a circular, cast-iron grate, which is ¾ in. thick and 6 ft. in
diameter and is set and secured horizontally in the pot. This grate
is perforated with holes ¾ in. in diameter, 2 in. apart, center to
center, and is similar to the Wetherill grate employed in zinc oxide
manufacture. The pot itself is about 2½ in. thick at the bottom,
thinning to about 1½ in. at the rim. It is supported on trunnions and
is geared for convenient turning by hand. The blast pipe which enters
the pot at the bottom is 6 in. in diameter.
Two roasting furnaces and six converters are rated nominally as a 90
ton plant. This rating is, however, considerably in excess of the
actual capacity, at least on certain ores. The time required for
desulphurization in the converter apparently depends a good deal upon
the character of the ore. The six converters may be arranged in a
single row, or in two rows of three in each. They are set so that the
rim of the pot, when upright, is about 12 ft. above the ground level.
A platform gives access to the pots. One man per shift can attend to
two pots. His work consists in charging them, which is done by gravity,
spreading out the charge evenly in the pot, closing any blow-holes
which may develop, and at the end of the operation raising the hood
(which covers the pot during the operation) and dumping the pot. The
work is easy. The conditions under which it is done are comfortable,
both as to temperature and atmosphere. Reports have shown a great
reduction in liability to lead-poisoning in the works where the H.-H.
process has been introduced.
A new charge is started by kindling a small wood or coal fire on the
grate, then throwing in a few shovelfuls of hot calcines, and finally
dropping in the regular charge of damp ore (plus the fluxes previously
referred to). The charge is introduced in stages, successive layers
being dropped in and spread out as the heat rises. At the beginning
the blast is very low—about 2 oz. It is increased as the hight of the
ore in the pot rises, finally attaining about 16 oz. The operation
goes on quietly, the smoke rising from the surface evenly and gently,
precisely as in a well-running blast furnace. While the charge is still
black on top, the hand can be held with perfect comfort, inside of
the hood, immediately over the ore. This explains, of course, why the
volatilization of silver and lead is insignificant. There is, moreover,
little or no loss of ore as dust, because the ore is introduced damp,
and the passage of the air through it is at low velocity. In the
interior of the charge, however, there is high temperature (evidently
much higher than has been stated in some descriptions), as will be
shown further on. The conditions in this respect appear to be analogous
to those of the blast furnace, which, though smelting at a temperature
of say 1200 deg. C. at the level of the tuyeres, suffers only a slight
loss of silver and lead by volatilization.
At the end of the operation in the H.-H. pot, the charge is dull red
at the top, with blow-holes, around which the ore is bright red.
Imperfectly worked charges show masses of well-fused ore surrounded
by masses of only partially altered ore, a condition which may be
ascribed to the irregular penetration of air through the charge,
affording good evidence of the important part which air plays in the
process. A properly worked charge is tipped out of the pot as a solid
cake, which in falling to the ground breaks into a few large pieces.
As they break, it appears that the interior of the charge is bright
red all through, and there is a little molten slag which runs out of
cavities, presumably spots where the chemical action has been most
intense. When cold, the thoroughly desulphurized material has the
appearance of slag-roasted galena. Prills of metallic lead are visible
in it, indicating reaction between lead sulphide and lead sulphate.
The columns of the structure supporting the pots should be of steel,
since fragments of the red-hot ore dumped on the ground are likely to
fall against them. To hasten the cooling of the ore, water is sometimes
played on it from a hose. This is bad, since some is likely to splash
into the still inverted pot, leading to cracks. The cracked pots at
certain works appear to be due chiefly to this cause, in the absence of
which the pots ought to last a long time, inasmuch as the conditions
to which they are subjected during the blowing process are not at all
severe. When the ore is sufficiently cold it is further broken up,
first by driving in wedges, and finally by sledging down to pieces
of orange size, or what is suitable for the blast furnace. These are
forked out, leaving the fine ore, which comes largely from the top of
the charge and is therefore only partially desulphurized. The fines
are, therefore, re-treated with a subsequent charge. The quantity is
not excessive; it may amount to 7 or 8 per cent. of the charge.
The breaking up of the desulphurized ore is one of the problems of the
process, the necessity being the reduction of several large pieces
of fused, or semi-fused, material weighing two or three tons each.
When done by hand only, as is usually (perhaps always) the practice,
the operation is rather expensive. It would appear, however, to
be not a difficult matter to devise some mechanical aids for this
process—perhaps to make it entirely mechanical. When done by hand, a
6-pot plant requires 6 men per shift sledging and forking. With 8-hour
shifts, this is 18 men for the breaking of about 60 tons of material,
which is about 3⅓ tons per man per 8 hours. With labor at 25c. per
hour, the cost of breaking the fused material comes to 60c. per ton. It
may be remarked, for comparison, that in breaking ore as it ordinarily
comes, coarse and fine together, a good workman would normally be
expected to break 5 to 5.5 tons in a shift of 8 hours.
The ordinary charge for the standard converter is about 8 tons (16,000
lb.) of an ore weighing 166 lb. per cu. ft. With a heavier ore, like
a high-grade galena, the charge would weigh proportionately more. The
time of working off a charge is decidedly variable. Accounts of the
operation of the process in Australia tell of charge-workings in 3
to 5 hours, but this does not correspond with the results reported
elsewhere, which specify times of 12 to 18 hours. Assuming an average
of 16 hours, which was the record of one plant, six converters would
have capacity for about 72 tons of charge per 24 hours, or about 58
tons of ore, the ratio of ore to flux being 4:1. The loss in weight
of the charge corresponds substantially to the replacement of sulphur
by oxygen, and the expulsion of carbon dioxide. The finished charge
contains on the average from 3 to 5 per cent. sulphur. This is
about the same as the result achieved in good practice in roasting
lead-bearing ores in hand-worked reverberatory furnaces, but curiously
the H.-H. product, in some cases at least, does not yield any matte,
to speak of, in the blast furnace; the product delivered to the latter
being evidently in such condition that the remaining sulphur is almost
completely burned off in the blast furnace. This is an important saving
effected by the process. In calculating the value of an ore, sulphur
is commonly debited at the rate of 25c. per unit, which represents
approximately the cost of handling and reworking the matte resulting
from it. The practically complete elimination of matte-fall rendered
possible by the H.-H. process may not be, however, an unmixed blessing.
There may be, for example, a small formation of lead sulphide which
causes trouble in the crucible and lead-well, and results in furnace
difficulties and the presentation of a vexatious between-product.
It may now be attempted to summarize the cost of the converting
process. Assuming the case of an ore assaying lead, 50 per cent.; iron,
15; sulphur, 22; silica, 8, and alumina, etc., 5, let it be supposed
that it is to be fluxed with pure limestone and pure quartz, with the
aim to make a slag containing silica, 30; ferrous oxide, 40; and lime,
20 per cent. A ton of ore will make, in round numbers, 1000 lb. of
slag, and will require 344 lb. of limestone and 130 lb. of quartz,
or we may say roughly one ton of flux must be added to four tons of
ore, wherefore the ore will constitute 80 per cent. of the charge. In
reducing the charge to 3 per cent. sulphur it will lose ultimately
through expulsion of sulphur and carbon dioxide (of the limestone)
about 20 per cent. in weight, wherefore the quantity of material to
be smelted in the blast furnace will be practically equivalent to
the raw sulphide ore in the charge for the roasting furnaces; but in
the roasting furnace the charge is likely to gain weight, because of
the formation of sulphates. Taking the charge, which I have assumed
above, and reckoning that as it comes from the roasting furnace it
will contain 10 per cent. sulphur, all in the form of sulphate, either
of lead or of lime, and that the iron be entirely converted to ferric
oxide, in spite of the expulsion of the carbon dioxide of the limestone
and the combustion of a portion of the sulphur of the ore as sulphur
dioxide, the charge will gain in weight in the ratio of 1:1.19. This,
however, is too high, inasmuch as a portion of the sulphur will remain
as sulphide while a portion of the iron may be as ferrous oxide. The
actual gain in weight will consequently be probably not more than
one-tenth. The following theoretical calculation will illustrate the
changes:
─────────────────────┬──────────────────────┬─────────────────────────
RAW CHARGE │ SEMI-ROASTED CHARGE │ FINISHED CHARGE
─────────────────────┼──────────────────────┼─────────────────────────
{1000 lb. Pb │ {1154 lb. PbO │ { 1154 lb. PbO
{ 300 lb. Fe │ { 428 lb. Fe₂O₃ │ { 428 lb. Fe₂O₃(?)
Ore { 160 lb. SiO₂ │ Ore { 160 lb. SiO₂ │ Ore { 160 lb. SiO₂
{ 100 lb. Al₂O₃,│ { 100 lb. Al₂O₃, │ { 100 lb. Al₂O₃,
etc. │ etc. │ etc.
{ 440 lb. S │ { 300 lb. S │ { 68 lb. S
│ │
{ 130 lb. SiO₂ │ { 130 lb. SiO₂ │ { 130 lb. SiO₂
Flux { 344 lb. CaCO₃ │ Flux { 193 lb. CaO │ Flux { 193 lb. CaO
│ 450 lb. O │
———— │ ———— │ ————
2474 lb. │ 2915 lb. │ 2233 lb.
│ │
│ 10% S. │ 3% S.
─────────────────────┴──────────────────────┴─────────────────────────
Ratios:
2474:2915 :: 1:1.18.
2915:2233 :: 1:0.76⅔.
2474:2233 :: 1:0.90.
It may be assumed that for every ton of charge (containing about 80 per
cent. of ore) there will be 1.1 ton of material to go to the converter,
and that the product of the latter will be 0.9 of the weight of the
original charge of raw material.
Each converter requires 400 cu. ft. of air per minute. The blast
pressure is variable, as different pots are always at different stages
of the process, but assuming the maximum of 16 oz. pressure, with a
blast main of sufficient diameter (at least 15 in.) and the blower
reasonably near the battery of pots, the total requirement is 21 h.p.
The cost of converting will be approximately as follows:
Labor, 3 foremen at $3.20 $ 9.60
“ 9 men at $2.50 22.50
Power, 21 h.p. at 30c 6.30
Supplies, repairs and renewals 5.00
——————
Total $43.40 = 60c. per ton of charge.
The cost of converting is, of course, reduced directly as the time is
reduced. The above estimate is based on unfavorable conditions as to
time required for working a charge.
The total cost of treatment, from the initial stage to the delivery of
the desulphurized ore to the blast furnaces, will be, per 2000 lb. of
charge, approximately as follows:
Crushing 1.0 ton at 10c $0.10
Mixing 1.0 ton at 10c .10
Roasting 1.0 ton at 63c .63
Delivering 1.1 ton to converters at 12c .13
Converting 1.1 ton at 60c .66
Breaking 0.9 ton at 60c .54
——-——-
Total $2.16
The cost per ton of ore will be 2.16 ÷ 0.80 = $2.70. Making allowance
for the crushing of the ore, which is not ordinarily included in the
cost of roasting, and possibly some overestimates, it appears that the
cost of desulphurization by this method, under the conditions assumed
in this paper, is rather higher than in good practice with ordinary
hand-worked furnaces, but it is evident that the cost can be reduced to
approximately the same figure by introduction of improvements, as for
example in breaking the desulphurized ore, and by shortening the time
of converting, which is possible in the case of favorable ores. The
chief advantage must be, however, in the further stage of the smelting.
As to this, there is the evidence that the Broken Hill Proprietary
Company was able to smelt the same quantity of ore in seven furnaces,
after the introduction of the Huntington-Heberlein process, that
formerly required thirteen. A similar experience is reported at
Friedrichshütte, Silesia.
This increase in the capacity of the blast furnace is due to three
things: (1) In delivering to the furnace a charge containing a reduced
percentage of fine ore, the speed of the furnace is increased, i.e.,
more tons of ore can be smelted per square foot of hearth area. (2)
There is less roasted matte to go into the charge. (3) Under some
conditions the percentage of lead in the charge can be increased,
reducing the quantity of gangue that must be fluxed.
It is difficult to generalize the economy that is effected in the
blast-furnace process, since this must necessarily vary within wide
limits because of the difference in conditions. An increase of 60 to
100 per cent. in blast-furnace capacity does not imply a corresponding
reduction in the cost of smelting. The fuel consumption per ton of ore
remains the same. There is a saving in the power requirements, because
the smelting can be done with a lower blast pressure; also, a saving
in the cost of reworking matte. There will, moreover, be a saving in
other labor, in so far as portions thereof are not already performed
at the minimum cost per ton. The net result under American conditions
of silver-lead smelting can only be determined closely by extensive
operations. That there will be an important saving, however, there is
no doubt.
The cost of smelting a ton of charge at Denver and Pueblo, exclusive
of roasting and general expense, is about $2.50, of which about $0.84
is for coke and $1.66 for labor, power and supplies. General expense
amounts to about $0.16 additional. If it should prove possible to
smelt in a given plant 50 per cent. more ore than at present without
increase in the total expense, except for coke, the saving per ton of
charge would be 70c. That is not to be expected, but the half of it
would be a satisfactory improvement. With respect to sulphur in the
charge, the cost is commonly reckoned at 25c. per unit. As compared
with a charge containing 2 per cent. of sulphur there would be a saving
rising toward 50c. per ton as the maximum. It is reasonable to reckon,
therefore, a possible saving of 75c. per ton of charge in silver-lead
smelting, no saving in the cost of roasting, and an increase of about
3 per cent. in the extraction of lead, and perhaps 1 per cent. in the
extraction of silver, as the net results of the application of the
Huntington-Heberlein process in American silver-lead smelting.
On a charge averaging 12 per cent. lead and 33 oz. silver per ton,
an increase of 3 per cent. in the extraction of lead and 1 per
cent. in the extraction of silver would correspond to 25c. and 35c.
respectively, reckoning lead at 3.5c. per lb., and silver at 60c. per
oz. In this, however, it is assumed that all lead-bearing ores will
be desulphurized by this process, which practically will hardly be
the case. A good deal of pyrites, containing only a little lead, will
doubtless continue to be roasted in Brückner cylinders, and other
mechanical furnaces, which are better adapted to the purpose than are
the lime-roasting pots. Moreover, a certain proportion of high-grade
lead ore, which is now smelted raw, will be desulphurized outside of
the furnace, at additional expense. It is comparatively simple to
estimate the probable benefit of the Huntington-Heberlein process in
the case of smelting works which treat principally a single class of
ore, but in such works as those in Colorado and Utah, which treat a
wide variety of ores, we must anticipate a combination process, and
await results of experience to determine just how it will work out.
It should be remarked, moreover, that my estimates do not take into
account the royalty on the process, which is an actual debit, whether
it be paid on a tonnage basis or be computed in the form of a lump sum
for the license to its use.
However, in view of the immense tonnage of ore smelted annually for
the extraction of silver and lead, it is evident that the invention of
lime-roasting by Huntington and Heberlein was an improvement of the
first order in the metallurgy of lead.
In the case of non-argentiferous galena, containing 65 per cent. of
lead (as in southeastern Missouri), comparison may be made with the
slag-roasting and blast-furnace smelting of the ore. Here, no saving
in cost of roasting may be reckoned and no gain in the speed of the
blast furnaces is to be anticipated. The only savings will be in
the increase in the extraction of lead from 92 to 98 per cent., and
the elimination of matte-roasting, which latter may be reckoned as
amounting to 50c. per ton of ore. The extent of the advantage over
the older method is so clearly apparent that it need not be computed
any further. In comparison with the Scotch-hearth bag-house method of
smelting, however, the advantage, if any, is not so certain. That
method already saves 98 per cent. of the lead, and on the whole is
probably as cheap in operation as the Huntington-Heberlein could be
under the same conditions. The Huntington-Heberlein method has replaced
the old roast-reaction method at Tarnowitz, Silesia, but the American
Scotch-hearth method as practised near St. Louis is likely to survive.
A more serious competitor will be, however, the Savelsberg process,
which appears to do all that the Huntington-Heberlein process does,
without the preliminary roasting. Indeed, if the latter be omitted
(together with its estimated expense of 63c. per ton of charge, or
79c. per ton of ore), all that has been said in this paper as to the
Huntington-Heberlein process may be construed as applying to the
Savelsberg. The charge is prepared in the same way, the method of
operating the converters is the same, and the results of the reactions
in the converters are the same. The litigation which is pending between
the two interests, Messrs. Huntington and Heberlein claiming that
Savelsberg infringes their patents, will be, however, a deterrent to
the extension of the Savelsberg process until that matter be settled.
The Carmichael-Bradford process may be dismissed with a few words. It
is similar to the Savelsberg, except that gypsum is used instead of
limestone. It is somewhat more expensive because the gypsum has to be
ground and calcined. The process works efficiently at Broken Hill,
but it can hardly be of general application, because gypsum is likely
to be too expensive, except in a few favored localities. The ability
to utilize the converter gases for the manufacture of sulphuric acid
will cut no great figure, save in exceptional cases, as at Broken
Hill, and anyway the gases of the other processes can be utilized for
the same purpose, which is in fact being done in connection with the
Huntington-Heberlein process in Silesia.
The cost of desulphurizing a ton of galena concentrate by the
Carmichael-Bradford process is estimated by the company controlling
the patents as follows, labor being reckoned at $1.80 per eight hours,
gypsum at $2.40 per 2240 lb., and coal at $8.40 per 2240 lb.:
0.25 ton of gypsum $0.60
Dehydrating and granulating gypsum .48
Drying mixture of ore and gypsum .12
Converting 0.24
Spalling sintered material .12
0.01 ton coal .08
——-——-
Total $1.64
The value of the lime in the sintered product is credited at 12c.,
making the net cost $1.52 per 2240 lb. of ore.
The cost allowed for converting may be explained by the more rapid
action that appears to be attained with the ores of Broken Hill than
with some ores that are treated in North America, but the low figure
estimated for spalling the sintered material appears to be highly
doubtful.
The theory of the lime-roasting processes is not yet well established.
It is recognized that the explanation offered by Huntington and
Heberlein in their original patent specification is erroneous. There is
no good evidence in their process, or any other, of the formation of
the higher oxide of lime, which they suggest.
At the present time there are two views. In one, formulated most
explicitly by Professor Borchers, there is formed in this process a
plumbate of calcium, which is an active oxidizing agent. A formation of
this substance was also described by Carmichael in his original patent,
but he considered it to be the final product, not the active oxidizing
agent.
In the other view, the lime, or limestone, serves merely as a diluent
of the charge, enabling the air to obtain access to the particles of
galena, without liquefaction of the latter. The oxidation of the lead
sulphide is therefore effected chiefly by the air, and the process
is analogous to what takes place in the bessemer converter or in the
Germot process of smelting, or perhaps more closely to what might
happen in an ordinary roasting furnace, provided with a porous hearth,
through which the air supply would be introduced. Roasting furnaces of
that design have been proposed, and in fact such a construction is now
being tested for blende roasting in Kansas.
Up to the present time, the evidence is surely too incomplete to enable
a definite conclusion to be reached. Some facts may, however, be stated.
There is clearly reaction to a certain extent between lead sulphide
and lead sulphate, as in the reverberatory smelting furnace, because
prills of metallic lead are to be observed in the lime-roasted charge.
There is a formation of sulphuric acid in the lime-roasting, upon the
oxidizing effect of which Savelsberg lays considerable stress, since
its action is to be observed on the iron work in which it condenses.
Calcium sulphate, which is present in all of the processes, being
specifically added in the Carmichael-Bradford, evidently plays an
important chemical part, because not only is the sulphur trioxide
expelled from the artificial gypsum, but also it is to a certain
extent expelled from the natural gypsum, which is added in the
Carmichael-Bradford process; in other words, more sulphur is given off
by the charge than is contained by the metallic sulphides alone.
Further evidence that lime does indeed play a chemical part in the
reaction is presented by the phenomena of lime-roasting in clay dishes
in the assay muffle, wherein the air is certainly not blown through the
charge, which is simply exposed to superficial oxidation as in ordinary
roasting.
The desulphurized charge dropped from the pot is certainly at much
below the temperature of fusion, even in the interior, but we have no
evidence of the precise temperature condition during the process itself.
Pyrite and even zinc blende in the ore are completely oxidized. This,
at least, indicates intense atmospheric action.
The papers by Borchers,[39] Doeltz,[40] Guillemain,[41] and
Hutchings[42] may profitably be studied in connection with the
reactions involved in lime-roasting. The conclusion will be, however,
that their precise nature has not yet been determined. In view of the
great interest that has been awakened by this new departure in the
metallurgy of lead, it is to be expected that much experimental work
will be devoted to it, which will throw light upon its principles, and
possibly develop it from a mere process of desulphurization into one
which will yield a final product in a single operation.
PART VI
OTHER METHODS OF SMELTING
THE BORMETTES METHOD OF LEAD AND COPPER SMELTING[43]
BY ALFREDO LOTTI
(September 30, 1905)
It is well known that, in order to obtain a proper fusion in lead
and copper ore-smelting, it is not only advantageous, but often
indispensable, that a suitable proportion of slag be added to the
charge. In the treatment of copper matte in the converter, the total
quantity of slag must be resmelted, inasmuch as it always retains a
notable quantity of the metal; while in the smelting of lead ore in the
blast furnace, the addition of slag is mainly intended to facilitate
the operation, avoiding the use of strong air pressure and thus
diminishing the loss of lead. The proportion of slag required sometimes
amounts to 30 to 35 per cent. of the weight of the ore.
Inasmuch as the slag is usually added in lump form, cold, its original
heat (about 400 calories per kilogram) is completely lost and an
intimate mixture with the charge cannot be obtained. For this reason, I
have studied the agglomeration of lead and copper ores with fused slag,
employing a variable proportion according to the nature of the ore
treated. In the majority of cases, and with some slight modifications
in each particular case, by incorporating the dry or slightly moistened
mineral with the predetermined quantity of liquid slag, and by rapidly
stirring the mixture so as to secure a proper subdivision of the slag
and the mineral, there is produced a spongy material, largely composed
of small pieces, together with a simultaneous evolution of dense fumes
of sulphur, sulphur dioxide, and sulphur trioxide. By submitting this
spongy material to an air blast, the sulphur of the mineral is burned,
the temperature rising in the interior of the mass to a clear red
heat. Copious fumes of sulphur dioxide and trioxide are given off,
and at times a yellowish vapor of sulphur, which condenses in drops,
especially if the ore is pyritous.
At the end of from one to three hours, according to the quantity of
sulphur contained in the material under treatment and the amount of the
air pressure, the desulphurization of the ore, so far as it has come
in contact with the air, is completed, and the mass, now thoroughly
agglomerated, forms a spongy but compact block. It is then only
necessary to break it up and smelt it with the requisite quantity of
flux and coke. The physical condition of the material is conducive to
a rapid and economical smelting, while the mixture of the sulphide,
sulphate and oxide leads to a favorable reaction in the furnace.
In employing this method, it sometimes happens that ores rich in
sulphur produce during the smelting a little more matte than when the
ordinary system of roasting is employed. In such instances, in order
to avoid or to diminish the cost of re-treatment of the matte, it is
best to agglomerate a portion thereof with the crude mineral and the
slag. This has the advantage of oxidizing the matte, which acts as a
ferruginous flux in the smelting.
The system described above leads to considerable economy, especially
in roasting, as the heat of the scoria, together with that given off
in the combustion of the sulphur, is almost always sufficient for the
agglomeration and desulphurization of the mineral; while, moreover, it
reduces the cost of smelting in the blast furnace. Although the primary
desulphurization is only partial (about 50 per cent.), it continues
in the blast furnace, since the mineral, agglomerated with the slag,
assumes a spongy form and thereby presents an increased surface to the
action of the air. The sulphur also acts as a fuel and does not produce
an excessive quantity of matte.
The system will prove especially useful in the treatment of
argentiferous lead ore, since, by avoiding the calcination in a
reverberatory furnace, loss of silver is diminished. It appears,
however, that, contrary to the reactions which occur in the
Huntington-Heberlein process, a calcareous or basic gangue is not
favorable to this process, if the proportion be too great.
The following comparison has been made in the case of an ore containing
62 to 65 per cent. of lead, 16 to 17 per cent. sulphur, 10 to 11 per
cent. zinc, 0.4 per cent. copper, and 0.222 per cent. silver, in which
connection it is to be remarked that, in general, the less zinc there
is in the ore the better are the results.
[Illustration: FIG. 21.—Elevation and Plan of Converting Chambers.]
_Ordinary Method._—Roast-reduction. Cost per 1000 kg. of crude ore:
1. Roasting in reverberatory furnace:
Labor $0.70
Fuel 1.50
Repairs and supplies .05
————- $2.25
2. Smelting in water-jacket:
Labor $1.01
Fuel 2.20
Repairs and supplies .03
Fluxes .50
————- 3.74
————-
Total $5.99
_Bormettes Method._—Agglomeration with slag, pneumatic desulphurization
and smelting in water-jacket:
1. Agglomeration and desulphurization:
Labor $0.42
Repairs and supplies 0.05
———- $0.47
2. Smelting in water-jacket:
Labor $0.90
Fuel 1.91
Repairs and supplies .03
Fluxes .42
————- 3.26
————-
Total $3.73
This shows a difference in favor of the new method of $2.26 per ton of
ore, without taking into account the savings realized by a much more
speedy handling of the operation, which would further reduce the cost
to approximately $2.50 per ton.
[Illustration: FIG. 22.—Details of Transfer Cars.]
In the above figures, no account has been taken of general expenses,
which per ton of ore are reduced because of the greater rapidity of the
process, enabling a larger quantity of ore to be smelted in a given
time. Making allowance for this, the saving will amount to an average
of $2.40 per 1000 kg., a figure which will naturally vary according
to the prices for fuel, labor, and the quantity of matte which it may
be necessary to re-treat. If the quantity of matte does not exceed
10 per cent. of the weight of the ore, it can be desulphurized by
admixture with the ore, without use of other fuel. If, however, the
proportion of matte rises to 20 parts per 100 parts of ore (a maximum
which ought not to be reached in good working), it is necessary to
roast a portion of it. Under unfavorable conditions, consequently,
the saving effected by this process may be reduced to $2 @ $2.20 per
1000 kg., and even to as little as $1.40 @ $1.60. The above reckonings
are, however, without taking any account of the higher extraction of
lead and silver, which is one of the great advantages of the Bormettes
process.
[Illustration: FIG. 23.—Latest Form of Converter. (Section on A B.)]
The technical results obtained in the smelting of an ore of the above
mentioned composition are as follows:
────────────────────────────────────┬─────────────┬─────────────
│ ORDINARY │ BORMETTES
│ METHOD │ METHOD
────────────────────────────────────┼─────────────┼─────────────
Coke, per cent. of the charge │ 14 │ 12
Blast pressure, water gage │12 to 20 cm. │12 to 14 cm.
Tons of charge smelted per 24 hr │ 20 │ 25
Tons of ore smelted per 24 hr │ 8 │ 10
Lead assay of slag │0.80 to 0.90%│0.20 to 0.40%
Matte-fall, per cent. of ore charged│ 5 to 10 │ 10 to 15
Lead extraction │ 90% │ 92%
Silver extraction │ 95% │ 98%
────────────────────────────────────┴─────────────┴─────────────
[Illustration: FIG. 24.—Latest Form of Converter. (Section on C D.)]
The higher extractions of lead and silver are explained by the fact
that the loss of metals in roasting is reduced, while, moreover, the
slags from the blast furnace are poorer than in the ordinary process
of smelting. The economy in coke results from the greater quantity of
sulphur which is utilized as fuel, and from the increased fusibility of
the charge for the blast furnace.
The new system of desulphurization enables the charge to be smelted
with a less quantity of fresh flux, by the employment in its place of a
greater proportion of foul slag. The reduction in the necessary amount
of flux is due not only to the increased fusibility of the agglomerated
charge, but principally to the fact that in this system the formation
of silicates of lead (which are produced abundantly in ordinary
slag-roasting) is almost nil. It is therefore unnecessary to employ
basic fluxes in order to reduce scorified lead.
[Illustration: FIG. 25.—Latest Form of Converter. (Plan.)]
The losses of metal in the desulphurization are less than in the
ordinary method, because the crude mineral remains only a short time
(from one to three hours) in the apparatus for desulphurization and
agglomeration, and the temperature of the process is lower. The
blast-furnace slags are poorer, because there is no formation of
silicate of lead during the agglomeration.
The Bormettes method, in so far as the treatment of lead ore is
concerned, may be considered a combination process of roast-reaction,
of roast-reduction, and of precipitation-smelting. It is not, however,
restricted to the treatment of lead ore. It may also be applied
to the smelting of pyritous copper-bearing ores. In an experiment
with cupriferous pyrites, containing 20 to 25 per cent. sulphur, it
succeeded in agglomerating and smelting them without use of any fuel
for calcination, effecting a perfect smelting, analogous to pyrite
smelting, with the production of a matte of sufficient degree of
concentration.
The first cost of plant installation is very much reduced by the
Bormettes method, inasmuch as the ordinary roasting furnaces are almost
entirely dispensed with, apparatus being substituted for them which
cost only one-third or one-fourth as much as ordinary furnaces. The
process presents the advantage, moreover, of being put into immediate
operation, without any expenditure of excess fuel.
The apparatus required in the process is illustrated in Figs. 21-25.
The apparatus for desulphurization and agglomeration consists of a
cast-iron box, composed of four vertical walls, of which two incline
slightly toward the front. These inclined walls carry the air-boxes.
The other two walls are formed, the one in front by the doors which
give access to the interior, and the other in the rear by a straight
plate. The whole arrangement is surmounted by a hood. The four pieces
when assembled form a box without bottom. Several of these boxes
are combined as a battery. The pots in which the agglomeration and
desulphurization are effected are moved into these boxes on suitable
cars, in the manner shown in the first engraving. A later and more
improved form is shown, however, in Figs. 23-25.
This process, which is the invention of A. Lotti and has been patented
in all the principal countries, is in successful use at the works of
the Société Anonyme des Mines de Bormettes, at Bormettes, La Londe
(Var), France. Negotiations are now in progress with respect to its
introduction elsewhere in Europe.
THE GERMOT PROCESS[44]
BY WALTER RENTON INGALLS
(November 1, 1902)
According to F. Laur, in the _Echo des Mines_ (these notes are
abstracted from _Oest. Zeit._, L., xl, 55, October 4, 1902), A. Germot,
of Clichy, France, made experiments some years ago upon the production
of white lead directly from galena. These led Catelin to attempt the
recovery of metallic lead in a similar way. If air be blown in proper
quantity into a fused mass of lead sulphide the following reaction
takes place:
2PbS + 2O = SO₂ + Pb + PbS.
Thus one-half of the lead is reduced, and it is found collects all the
silver of the ore; the other half is sublimed as lead sulphide, which
is free from silver. The reaction is exothermic to the extent that
the burning of one-half the sulphur of a charge should theoretically
develop sufficient heat to volatilize half of the charge and smelt the
other half. This is almost done in practice with very rich galena,
but not so with poorer ore. The temperature of the furnace must be
maintained at about 1100 deg. C. throughout the whole operation, and
there are the usual losses of heat by radiation, absorption by the
nitrogen of the air, etc. Deficiencies in heat are supplied by burning
some of the ore to white lead, which is mixed with the black fume
(PbS) and by the well-known reactions reduced to metal with evolution
of sulphur dioxide. The final result is therefore the production of
(1) pig lead enriched in silver; (2) pig lead free from silver; (3) a
leady slag; and (4) sulphur dioxide. In the case of ores containing
less than 75 per cent. Pb the gangue forms first a little skin and
then a thick hard crust which soon interferes with the operation,
especially if the ore be zinkiferous. This difficulty is overcome by
increasing the temperature or by fluxing the ore so as to produce a
fusible slag. A leady slag is always easily produced; this is the only
by-product of the process. The theoretical reaction requires 600 cu. m.
of air, assuming a delivery of 50 per cent. from the blower, and at one
atmosphere pressure involves the expenditure of 18 h.p. per 1000 kg. of
galena per hour.
[Illustration: FIG. 26.—Plan and Elevation of Smelting Plant at Clichy.]
The arrangement of the plant at Clichy is shown diagrammatically in
Fig. 26. There is a round shaft furnace, 0.54 meter in diameter and
4.5 meters high. Power is supplied to the blower C through the pulley
G and the shaft DD. The compressed air is accumulated in the reservoir
R, whence it is conducted by the pipe to the tuyere which is suspended
inside of the furnace by means of a chain, whereby it can be raised
or lowered. O₁ and O₂ are tap-holes. L is a door and N an
observation tube. A is the charge tube. X is the pipe which conveys the
gas and fume to the condensation chambers. T is the pipe through which
the waste gases are drawn. V is the exhauster and S is the chimney.
K₁ and K₂ are tilting crucible furnaces for melting lead and
galena.
After the furnace has been properly heated, 100 kg. of lead melted in
K₁ are poured in through the cast-iron pipe P, and after that about
200 kg. of pure, thoroughly melted galena from K₂. Ore containing 70
to 80 per cent. Pb must be used for this purpose. The blast of air is
then introduced into the molten galena, and from 1000 to 3000 kg. of
ore is gradually charged in through the tube A. During this operation
black fume (PbS) collects in the condensation chamber. All outlets are
closed against the external air. If the air blast is properly adjusted,
nothing but black fume is produced; if it begins to become light
colored, charging is discontinued and the blast of air is shut off.
Lead is then tapped through O₂, which is about 0.2 meter above the
hearth, so there is always a bath of lead in the bottom of the furnace;
but it is advisable now and then to tap off some through O₁, so as
gradually to heat up the bottom of the furnace. Hearth accretions are
also removed through O₁. The lead is tapped off through O₂ until
matte appears. The tap hole is then closed, the tuyere is lowered and
the blast is turned into the lead in order to oxidize it and completely
desulphurize the sulphur combinations, which is quickly done. The
oxide of lead is scorified as a very fusible slag, which is tapped off
through O₂, and more ore is then charged in upon the lead bath and
the cycle of operations is begun again.
PART VII
DUST AND FUME RECOVERY
FLUES, CHAMBERS AND BAG-HOUSES
DUST CHAMBER DESIGN
BY MAX J. WELCH
(September 1, 1904)
Only a few years ago smelting companies began to recognize the
advantage of large chambers for collecting flue dust and condensing
fumes. The object is threefold: First, profit; second, to prevent law
suits with surrounding agricultural interests; third, cleanliness about
the plant. It is my object at present to discuss the materials used in
construction and general types of cross-section.
Most of the old types of chambers are built after one general pattern,
namely, brick or stone side walls and arch roof, with iron buckstays
and tie rods. The above type is now nearly out of use, because it is
short-lived, expensive, and dangerous to repair, while the steel and
masonry are not used to good advantage in strength of cross-section.
With the introduction of concrete and expanded metal began a new
era of dust-chamber construction. It was found that a skeleton of
steel with cement plaster is very strong, light and cheap. The first
flue of the type shown in Fig. 29 was built after the design of E.
H. Messiter, at the Arkansas Valley smelter in Colorado. This flue
was in commission several years, conveying sulphurous gases from the
reverberatory roaster plant. The same company decided, in 1900, to
enlarge and entirely rebuild its dust-chamber system, and three types
of cross-section were adopted to meet the various conditions. All three
types were of cement and steel construction.
The first type, shown in Fig. 27, is placed directly behind the blast
furnaces. The cross-section is 273 sq. ft. area, being designed for a
10-furnace lead smelter. The back part is formed upon the slope of the
hillside and paved with 2.5 in. of brick. The front part is of ribbed
cast-iron plates. Ninety per cent. of the flue dust is collected in
this chamber and is removed, through sliding doors, into tram cars.
There is a little knack in designing a door to retain flue dust. It is
simply to make the bottom sill of the door frame horizontal for a space
of about 1 in. outside of the door slide.
The front part of the chamber, Fig. 27, is of expanded metal and
cement. The top is of 20 in. I-beams, spanning a distance of 24 ft.
with 15 in. cross-beams and 3 in. of concrete floor resting upon
the bottom flanges of the beams. This heavy construction forms the
foundation for the charging floor, bins, scales, etc.
[Illustration: FIG. 27.—Rectangular form of Concrete Dust Chamber.]
While dwelling upon this type of construction I wish to mention a
most important point, that of the proper factor of safety. Flue dust,
collected near the blast furnace, weighs from 80 to 100 lb. per cubic
foot, and the steel supports should be designed for 16,000 lb. extreme
fiber stress, when the chamber is three-quarters full of dust. If the
dust is allowed to accumulate beyond this point, the steel, being well
designed, should not be overstrained. Discussions as to strains in
bins have been aired by the engineering profession, but the present
question is “Where is a dust chamber a bin?” Experience shows that bin
construction should be adopted behind, or in close proximity to, the
blast furnaces.
Fig. 28 shows the second type of hopper-bottom flue adopted. It is
of very light construction, of 274 sq. ft. area in the clear. The
beginning of this flue being 473 ft. from the blast furnaces removes
all possibility of any material floor-load, as the dust is light in
weight and does not collect in large quantities. The hopper-bottom
floor is formed of 4 in. concrete slabs, in panels, placed between 4
in. I-beams. Cast-iron door frames, with openings 12 × 16 in., are
placed on 5 ft. centers. The concrete floor is tamped in place around
the frames. The side walls and roof are built of 1 in. angles, expanded
metal, and plastered to 2.5 in. thickness. At every 10 ft. distance,
pilaster ribs built of 2 in. angles, latticed and plastered, form the
wind-bracing and arch roof support.
[Illustration: FIG. 28.—Arched form of Concrete Dust Chamber.]
Fig. 29 shows the beehive construction. This chamber is of 253 sq. ft.
cross-sectional area. It is built of 2 in. channels, placed 16 in.
centers, tied with 1 × 0.125 in. steel strips. The object of the strips
is to support the 2 in. channels during erection. No. 27 gage expanded
metal lath was wired to the inside of the channels and the whole
plastered to a thickness of 3 in. The inside coat was plastered first
with portland cement and sand, one to three, with about 5 per cent.
lime. The filling between ribs is one to four, and the outside coat one
to three.
The above types of dust chamber have been in use over three years at
Leadville. Cement and concrete, in conjunction with steel, have been
used in Utah, Montana and Arizona, in various types of cross-section.
The results show clearly where not to use cement; namely, where
condensing sulphur fumes come in contact with the walls, or where
moisture collects, forming sulphuric acid. The reason is that portland
cement and lime mortar contain calcium hydrate, which takes up sulphur
from the fumes, forming calcium sulphate. In condensing chambers, this
calcium sulphate takes up water, forming gypsum, which expands and
peels off.
[Illustration: FIG. 29.—Beehive form of Concrete Dust Chamber.]
In materials of construction it is rather difficult to get something
that will stand the action of sulphur fumes perfectly. The lime mortar
joints in the old types of brick flues are soon eaten away. The arches
become weak and fall down. I noted a sheet steel condensing system,
where in one year the No. 12 steel was nearly eaten through. With a
view of profiting by past experience, let us consider the acid-proof
materials of construction, namely, brick, adobe mortar, fire-clay, and
acid-proof paint. Also, let us consider at what place in a dust-chamber
system are we to take the proper precaution in the use of these
materials.
At smelting plants, both copper and lead, it is found that near the
blast furnaces the gases remain hot and dry, so that concrete, brick
or stone, or steel, can safely be used. Lead-blast furnace gases will
not injure such construction at a distance of 6 or 8 ft. away from the
furnaces. For copper furnaces, roasters or pyritic smelting, concrete
or lime mortar construction should be limited to within 200 or 300 ft.
of the furnaces.
Another type of settling chamber is 20 ft. square in the clear, with
concrete floor between beams and steel hopper bottom. This chamber
is built within 150 ft. distance from the blast furnaces, and is one
of the types used at the Shannon Copper Company’s plant at Clifton,
Arizona. After passing the 200 ft. mark, there is no need of expensive
hopper design. The amount of flue dust settled beyond this point is so
small that it is a better investment to provide only small side doors
through which the dust can be removed. The ideal arrangement is to have
a system of condensing chambers, so separated by dampers that either
set can be thrown out for a short time for cleaning purposes, and the
whole system can be thrown in for best efficiency.
As to cross-section for condensing chambers, I consider that the
following will come near to meeting the requirements. One, four, and
six, concrete foundation; tile drainage; 9 in. brick walls, laid in
adobe mortar, pointed on the outside with lime mortar; occasional
strips of expanded metal flooring laid in joints; the necessary
pilasters to take care of the size of cross-section adopted; the top
covered with unpainted corrugated iron, over which is tamped a concrete
roof, nearly flat; concrete to contain corrugated bars in accordance
with light floor construction; and lastly, the corrugated iron to have
two coats of graphite paint on under side.
The above type of roof is used under slightly different conditions over
the immense dust chamber of the new Copper Queen smelter at Douglas,
Arizona. The paint is an important consideration. Steel work imbedded
in concrete should never be painted, but all steel exposed to fumes
should be covered by graphite paint. Tests made by the United States
Graphite Company show that for stack work the paint, when exposed to
acid gases, under as high a temperature as 700 deg. F., will wear well.
CONCRETE IN METALLURGICAL CONSTRUCTION[45]
BY HENRY W. EDWARDS
The construction of concrete flues of the section shown in Fig. 31
gives better results than that shown in Fig. 30, being less liable
to collapse. It costs somewhat more to build owing to the greater
complication of the crib, which, in both cases, consists of an interior
core only. For work 4 in. in thickness and under, I recommend the
use of rock or slag crushed to pass through a 1.5 in. ring. Although
concrete is not very refractory, it will easily withstand the heat
of the gases from a set of ordinary lead-or copper-smelting blast
furnaces, or from a battery of calcining or roasting furnaces. I have
never noticed that it is attacked in any way by sulphur dioxide or
other furnace gas.
[Illustration: FIGS. 30 and 31.—Sections of Concrete Flues.]
Shapes the most complicated to suit all tastes in dust chambers can
be constructed of concrete. The least suitable design, so far as the
construction itself is concerned, is a long, wide, straight-walled,
empty chamber, which is apt to collapse, either inwards or outwards,
and, although the outward movement can be prevented by a system of
light buckstays and tie-rods, the tendency to collapse inwards is
not so simply controlled in the absence of transverse baffle walls.
The tendency, so far as the collection of mechanical flue dust is
concerned, appears to be towards a large empty chamber, without
baffles, etc., in which the velocity of the air currents is reduced to
a minimum, and the dust allowed to settle. In the absence of transverse
baffle walls to counteract the collapsing tendency, it seems best to
design the chamber with a number of stout concrete columns at suitable
intervals along the side and end walls—the walls themselves being made
only a few inches thick with woven-wire screen or “expanded metal”
buried within them. The wire skeleton should also be embedded into the
columns in order to prevent the separation of wall and the columns.
This method of constructing is one that I have followed with very
satisfactory results as far as the construction itself is concerned.
[Illustration: FIG. 32.—Concrete Dust Chamber at the Guillermo Smelting
Works, Palomares, Spain. (Horizontal section.)]
Figs. 32 and 33 show a chamber designed and erected at the Don
Guillermo Smelting Works at Palomares, Province of Murcia, Spain.
Figs. 34 and 35 show a design for the smelter at Murray Mine, Sudbury,
Ontario, in which the columns are hollow, thus economizing concrete
material. For work of this kind the columns are built first and the
wire netting stretched from column to column and partly buried within
them. The crib is then built on each side of the netting, a gang of men
working from both sides, and is built up a yard or so at a time as the
work progresses. Doors of good size should be provided for entrance
into the chamber, and as they will seldom be opened there is no need
for expensive fastenings or hinges.
[Illustration: FIG. 33.—Concrete Dust Chamber at the Guillermo Smelting
Works, Palomares, Spain. (End elevation.)]
_Foundations for Dynamos and other Electrical Machinery._—Dry concrete
is a poor conductor of electricity, but when wet it becomes a fairly
good conductor. Therefore, if it be necessary to insulate the
electrical apparatus, the concrete should be covered with a layer of
asphalt.
[Illustration: FIG. 34.—Concrete Dust Chamber designed for smelter at
Murray Mine, Sudbury, Ontario, Can. There are eight 9 ft. sections in
the plan.]
_Chimney Bases._—Fig. 36 shows the base for the 90 ft. brick-stack at
Don Guillermo. The resemblance to masonry is given by nailing strips of
wood on the inside of the crib.
[Illustration: FIG. 35.—Concrete Dust Chamber designed for smelter at
Murray Mine, Sudbury, Ontario, Can. (End elevation.)]
_Retaining-Walls._—Figs. 37, 38, and 39 show three different styles
of retaining-walls, according to location. These walls are shown
in section only, and show the placing of the iron reenforcements.
Retaining-walls are best built in panels (each panel being a day’s
work), for the reason that horizontal joints in the concrete are
thereby avoided. The alternate panels should be built first and the
intermediate spaces filled in afterward. Should there be water behind
the wall it is best to insert a few small pipes through the wall, in
order to carry it off; this precaution is particularly important in
places where the natural surface of the ground meets the wall, as
shown in Figs. 37 and 38. If a wooden building is to be erected on the
retaining-wall, it is best to bury a few 0.75 in. bolts vertically in
the top of the wall, by which a wooden coping may be secured (see Figs.
37, 38, and 39), which forms a good commencement for the carpenter work.
[Illustration: FIG. 36.—Concrete Base for a 90 ft. Chimney at the
Guillermo Smelting Works, Palomares, Spain.]
Minimum thickness for a retaining-wall, having a liberal quantity
of iron embedded therein, is 20 in. at the bottom and 10 in. at the
top, with the taper preferably on the inner face. In the absence of
interior strengthening irons the thickness of the wall at the bottom
should never be less than one-fourth the total hight, and at the top
one-seventh of the hight; unless very liberal iron bracing be used,
the dimensions can hardly be reduced to less than one-seventh and
one-tenth respectively. Unbraced retaining-walls are more stable with
the batter on the outer face. Dry clay is the most treacherous material
that can be had behind a retaining-wall, especially if it be beaten
in, for the reason that it is so prone to absorb moisture and swell,
causing an enormous side thrust against the wall. When this material is
to be retained it is best to build the wall superabundantly strong—a
precaution which applies even to a dry climate, because the bursting
of a water-pipe may cause the damage. In order to avoid horizontal
joints it is best, wherever practicable, to build the crib-work in its
entirety before starting the concrete. In a retaining-wall 3 ft. thick
by 16 ft. high this is not practicable. The supporting posts and struts
can, however, be completed and the boards laid in as the wall grows,
in order not to interrupt the regular progress of the tamping. A good
finish may be produced on the exposed face of the wall by a few strokes
of the shovel up and down with its back against the crib.
[Illustration: FIGS. 37, 38, and 39.—Retaining-Walls of Concrete.]
In conclusion I wish to state that this paper is not written for the
instruction of the civil engineer, or for those who have special
experience in this line; but rather for the mining engineer or
metallurgist whose training is not very deep in this direction, and who
is so often thrown upon his own resources in the wilderness, and who
might be glad of a few practical suggestions from one who has been in a
like predicament.
CONCRETE FLUES[46]
BY EDWIN H. MESSITER
(September, 1904)
Under the heading “Flues,” Mr. Edwards refers to the Beehive
construction, a cross-section of which is shown in Fig. 31 of his
paper. A flue similar to this was designed by me about six years
ago,[47] and in which the walls, though much thinner than those
described by Mr. Edwards, gave entire satisfaction. These walls, from
2.25 in. thick throughout in the smaller flues to 3.25 in. in the
larger, were built by plastering the cement mortar on expanded-metal
lath, without the use of any forms or cribs whatever, at a cost of
labor generally less than $1 per sq. yd. of wall. Of course, where
plasterers cannot be obtained on reasonable terms, the cement can be
molded between wooden forms, though it is difficult to see how it can
be done with an interior core only, as stated by Mr. Edwards.
In regard to the effect of sulphur dioxide and furnace gases on the
cement, I have found that in certain cases this is a matter which
must be given very careful attention. Where there is sufficient heat
to prevent the existence of condensed moisture inside of the flue,
there is apparently no action whatever on the cement, but if the
concrete is wet, it is rapidly rotted by these gases. At points near
the furnaces there is generally sufficient heat not only to prevent
internal condensation of the aqueous vapor always present in the gases,
but also to evaporate water from rain or snow falling on the outside
of the flue. Further along a point is reached where rain-water will
percolate through minute cracks caused by expansion and contraction,
and reach the interior even though internal condensation does not occur
there in dry weather. From this point to the end of the flue the roof
must be coated on the outside with asphalt paint or other impervious
material. In very long flues a point may be reached where moisture will
condense on the inside of the walls in cold weather. From this point
to the end of the flue it is essential to protect the interior with an
acid-resisting paint, of which two or more coats will be necessary.
For the first coat a material containing little or no linseed oil is
best, as I am informed that the lime in the cement attacks the oil. For
this purpose I have used ebonite varnish, and for the succeeding coats
durable metal-coating. The first coat will require about 1 gal. of
material for each 100 sq. ft. of surface.
In one of the earliest long flues built of cement in this country, a
small part near the chimney was damaged as a result of failure to apply
the protective coating, the necessity for it not having been recognized
at the time of its construction. It may be said, in passing, that other
long brick flues built prior to that time were just as badly attacked
at points remote from the furnaces. In order to reduce the amount of
flue subject to condensation, the plastered flues have been built with
double lath having an intervening air-space in the middle of the wall.
In building thin walls of cement, such as flue walls, it is
particularly important to prevent them from drying before the cement
has combined with all the water it needs. For this reason the work
should be sprinkled freely until the cement is fully set. Much work of
this class has been ruined through ignorance by fires built near the
walls in cold weather, which caused the mortar to shell off in a short
time.
The great saving in cost of construction, which the concrete-steel flue
makes possible, will doubtless cause it to supersede other types to
even a greater extent than it has already done. If properly designed
this type of construction reduces the cost of flues by about one-half.
Moreover, the concrete-steel flue is a tight flue as compared with
one built of brick. There is a serious leakage through the walls of
the brick flues which is not easily observed in flues under suction
as most flues are, but when a brick flue is under pressure from a
fan the leakage is surprisingly apparent. In flues operating by
chimney-draft the entrance of cold air must cause a considerable loss
in the efficiency of the chimney, a disadvantage which would largely be
obviated by the use of the concrete-steel flue.
CONCRETE FLUES[48]
BY FRANCIS T. HAVARD
In discussion of Mr. Edwards’s interesting and valuable paper, I
beg to submit the following notes concerning the advantages and
disadvantages of the concrete flues and stacks at the plant of the
Anhaltische Blei-und Silber-werke. The flues and smaller stacks at the
works were constructed of concrete consisting generally of one part of
cement to seven parts of sand and jig-tailings but, in the case of the
under-mentioned metal concrete slabs, of one part of cement to four
parts of sand and tailings. The cost of constructing the concrete flue
approximated 5 marks per sq. m. of area (equivalent to $0.11 per sq.
ft.).
_Effect of Heat._—A temperature above 100 deg. C. caused the concrete
to crack destructively. Neutral furnace gases at 120 deg. C., passing
through an independent concrete flue and stack, caused so much damage
by the formation of cracks that, after two years of use, the stack,
constructed of pipes 4 in. thick, required thorough repairing and
auxiliary ties for every foot of hight.
_Effect of Flue Gases and Moisture._—The sides of the main flue, made
of blocks of 6 in. hollow wall-sections, 100 cm. by 50 cm. in area,
were covered with 2 in. or 1 in. slabs of metal concrete. In cases
where the flue was protected on the outside by a wooden or tiled roof,
and inside by an acid-proof paint, consisting of water-glass and
asbestos, the concrete has not been appreciably affected. In another
case, where the protective cover, both inside and outside, was of
asphalt only, the concrete was badly corroded and cracked at the end
of three years. In a third case, in which the concrete was unprotected
from both atmospheric influence on the outside, and furnace gases on
the inside, the flue was quite destroyed at the end of three years.
That portion of the protected concrete flue, near the main stack, which
came in contact only with dry, cold gases was not affected at all.
Gases alone, such as sulphur dioxide, sulphur trioxide, and others,
do not affect concrete; neither is the usual quantity of moisture
in furnace gases sufficient to damage concrete; but should moisture
penetrate from the outside of the flue, and, meeting gaseous SO₂ or
SO₃, form hydrous acids, then the concrete will be corroded.
_Effect of the Atmosphere Alone._—For outside construction work,
foundations and other structures not exposed to heat, moist acid gases
and chemicals, the concrete has maintained its reputation for cheapness
and durability.
_Effect of Crystallization of Contained Salts._—In chemical works,
floors constructed of concrete are sometimes unsatisfactory, for the
reason that soluble salts, noticeably zinc sulphate, will penetrate
into the floor and, by crystallizing in narrow confines, cause the
concrete to crack and the floor to rise in places.
BAG-HOUSES FOR SAVING FUME
BY WALTER RENTON INGALLS
(July 15, 1905)
One of the most efficient methods of saving fume and very fine dust in
metallurgical practice is by filtration through cloth. This idea is by
no means a new one, having been proposed by Dr. Percy, in his treatise
on lead, page 449, but he makes no mention of any attempt to apply it.
Its first practical application was found in the manufacture of zinc
oxide direct from ores, initially tried by Richard and Samuel T. Jones
in 1850, and in 1851 modified by Samuel Wetherill into the process
which continues in use at the present time in about the same form as
originally. In 1878 a similar process for the manufacture of white
lead direct from galena was introduced at Joplin, Mo., by G. T. Lewis
and Eyre O. Bartlett, the latter of whom had previously been engaged
in the manufacture of zinc oxide in the East, from which he obtained
his idea of the similar manufacture of white lead. The difference
in the character of the ore and other conditions, however, made it
necessary to introduce numerous modifications before the process became
successful. The eventual success of the process led to its application
for filtration of the fume from the blast furnaces at the works of the
Globe Smelting and Refining Company, at Denver, Colo., and later on for
the filtration of the fume from the Scotch hearths employed for the
smelting of galena in the vicinity of St. Louis.
In connection with the smelting of high-grade galena in Scotch hearths,
the bag-house is now a standard accessory. It has received also
considerable application in connection with silver-lead blast-furnace
smelting and in the desilverizing refineries. Its field of usefulness
is limited only by the character of the gas to be filtered, it being
a prerequisite that the gas contain no constituent that will quickly
destroy the fabric of which the bags are made. Bags are also employed
successfully for the collection of dust in cyanide mills, and other
works in which fine crushing is practised, for example, in the
magnetic separating works of the New Jersey Zinc Company, Franklin,
N. J. , where the outlets of the Edison driers, through which the ore
is passed, communicate with bag-filtering machines, in which the bags
are caused to revolve for the purpose of mechanical discharge. The
filtration of such dust is more troublesome than the filtration of
furnace fume, because the condensation of moisture causes the bags to
become soggy.
[Illustration: FIG. 40.—Bag-house, Globe Smelting Works.]
The standard bag-house employed in connection with furnace work is a
large room, in which the bags hang vertically, being suspended from
the top. The bags are simply tubes of cotton or woolen (flannel)
cloth, from 18 to 20 in. in diameter, and 20 to 35 ft. in length, most
commonly about 30 ft. In the manufacture of zinc oxide, the fume-laden
gas is conducted into the house through sheet-iron pipes, with suitably
arranged branches, from nipples on which the bags are suspended, the
lower end of the bag being simply tied up until it is necessary to
discharge the filtered fume by shaking. In the bag-houses employed in
the metallurgy of lead, the fume is introduced at the bottom into brick
chambers, which are covered with sheet-iron plates, provided with the
necessary nipples; or else into hopper-bottom, sheet-iron flues, with
the necessary nipples on top. In either case the bags are tied to the
nipples, and are tied up tight at the top, where they are suspended.
When the fume is dislodged by shaking the bags, it falls into the
chamber or hopper at the bottom, whence it is periodically removed.
The cost of attending a bag-house, collecting the fume, etc., varies
from about 10c. per ton of ore smelted in a large plant like the Globe,
to about 25c. per ton in a Scotch-hearth plant treating 25 tons of ore
per 24 hours.
No definite rules for the proportioning of filtering area to the
quantity of ore treated have been formulated. The correct proportion
must necessarily vary according to the volume of gaseous products
developed in the smelting of a ton of ore, the percentage of dust and
fume contained, and the frequency with which the bags are shaken.
It would appear, however, that in blast furnaces and Scotch-hearth
smelting a ratio of 1000 sq. ft. per ton of ore would be sufficient
under ordinary conditions. The bag-house originally constructed at
the Globe works had about 250 sq. ft. of filtering area per ton of
charge smelted, but this was subsequently increased, and Dr. Iles,
in his treatise on lead-smelting, recommends an equipment which would
correspond to about 750 sq. ft. per ton of charge. At the Omaha works,
where the Brown-De Camp system was used, there was 80,000 sq. ft. of
cloth for 10 furnaces 42 × 120 in., according to Hofman’s “Metallurgy
of Lead,” which would give about 1000 sq. ft. per ton of charge
smelted, assuming an average of eight furnaces to be in blast. A
bag-house in a Scotch-hearth smeltery, at St. Louis, had approximately
900 sq. ft. per ton of ore smelted. At the Lone Elm works, at Joplin,
the ratio was about 3500 sq. ft. per ton of ore smelted, when the
works were run at their maximum capacity. In the manufacture of zinc
oxide the bag area used to be from 150 to 200 sq. ft. per square foot
of grate on which the ore is burned, but at Palmerton, Pa. (the most
modern plant), the ratio is only 100:1. This corresponds to about 1400
sq. ft. of bag area per 2000 lb. of charge worked on the grate. In the
manufacture of zinc-lead white at Cañon City, Colo., the ratio between
bag area and grate area is 150:1.
Assuming the gas to be free, or nearly free, from sulphurous fumes, the
bags are made of unbleached muslin, varying in weight from 0.4 to 0.7
oz. avoirdupois per square foot. The cloth should have 42 to 48 threads
per linear inch in the warp and the same number in the woof. A kind of
cloth commonly used in good practice weighs 0.6 oz. per square foot and
has 46 threads per linear inch in both the warp and the woof.
The bags should be 18 to 20 in. in diameter. Therefore the cloth should
be of such width as to make that diameter with only one seam, allowing
for the lap. Cloth 62 in. in width is most convenient. It costs 4 to
5c. per yard. The seam is made by lapping the edges about 1 in., or
by turning over the edges and then lapping, in the latter case the
stitches passing through four thicknesses of the cloth. It should be
sewed with No. 50 linen thread, making two rows of double lock-stitches.
The thimbles to which the bags are fastened should be of No. 10 sheet
steel, the rim being formed by turning over a ring of 0.25 in. wire.
The bags are tied on with 2 in. strips of muslin. The nipples are
conveniently spaced 27 in. apart, center to center, on the main pipe.
The gas is best introduced at a temperature of 250 deg. F. Too high
a temperature is liable to cause them to ignite. They are safe at 300
deg. F., but the temperature should not be allowed to exceed that point.
The gas is cooled by passage through iron pipes of suitable radiating
surface, but the temperature should be controlled by a dial thermometer
close to the bag-house, which should be observed at least hourly, and
there should be an inlet into the pipe from the outside, so that, in
event of rise of temperature above 300 deg., sufficient cold air may be
admitted to reduce it within the safety limit.
In the case of gas containing much sulphur dioxide, and especially any
appreciable quantity of the trioxide, the bags should be of unwashed
wool. Such gas will soon destroy cotton, but wool with the natural
grease of the sheep still in it is not much affected. The gas from
Scotch hearths and lead-blast furnaces can be successfully filtered,
but the gas from roasting furnaces contains too much sulphur trioxide
to be filtered at all, bags of any kind being rapidly destroyed.
PART VIII
BLOWERS AND BLOWING ENGINES
ROTARY BLOWERS VS. BLOWING ENGINES FOR LEAD SMELTING
(April 27, 1901)
A note in the communication from S. E. Bretherton on “Pyritic Smelting
and Hot Blast,” published in the _Engineering and Mining Journal_
of April 13, 1901, refers to a subject of great interest to lead
smelters. Mr. Bretherton remarked that he had been recently informed
by August Raht that by actual experiment the loss with the ordinary
rotary blowers was 100 per cent. under 10 lb. pressure; that is, it
was possible to shut all the gates so that there was no outlet for the
blast to escape from the blower and the pressure was only 10 lb., or in
other words the blower would deliver no air against 10 lb. pressure.
For that reason Mr. Raht expressed himself as being in favor of blowing
engines for lead blast furnaces. This is of special interest, inasmuch
as it comes from one who is recognized as standing in the first rank of
lead-smelting engineers. Mr. Raht is not alone in holding the opinion
he does.
The rotary blower did good service in the old days when the air was
blown into the lead blast furnace at comparatively moderate pressure.
At the present time, when the blast pressure employed is commonly
40 oz. at least, and sometimes as high as 48 oz., the deficiencies
of the rotary blower have become more apparent. Notwithstanding the
excellent workmanship which is put into them by their manufacturers,
the extensive surfaces of contact are inherent to the type, and
leakage of air backward is inevitable and important at the pressures
now prevailing. The impellers of a rotary blower should not touch
each other nor the cylinders in which they revolve, but they are made
with as little clearance as possible, the surfaces being coated with
grease, which fills the clearance space and forms a packing. This
will not, however, entirely prevent leakage, which will naturally
increase with the pressure. Even the manufacturers of rotary blowers
admit the defects of the type, and concede that for pressures of 5
lb. and upward the cylinder blowing engine is the more economical.
Metallurgists are coming generally to the opinion, however, that
blowing engines are probably more economical for pressures of 4 lb. or
thereabouts, and some go even further. With the blowing engines the
air-joints of piston and cylinder are those of actual contact, and
the metallurgist may count on his cubic feet of air, whatever be the
pressure. Blowing engines were actually introduced several years ago
by M. W. Iles at what is now the Globe plant of the American Smelting
and Refining Company, and we believe their performance has been found
satisfactory.
The fancied drawback to the use of blowing engines is their greater
first cost, but H. A. Vezin, a mechanical engineer whose opinions carry
great weight, pointed out five years ago in the _Transactions_ of the
American Institute of Mining Engineers (Vol. XXVI) that per cubic
foot of air delivered the blowing engine was probably no more costly
than the rotary blower, but on the contrary cheaper, stating that the
first cost of a cylinder blower is only 20 to 25 per cent. more than
that of a rotary blower of the same nominal capacity and the engine
to drive it. The capacity of a rotary blower is commonly given as the
displacement of the impellers per revolution, without allowance for
slip or leakage backward. Mr. Vezin expressed the opinion that for the
same actual capacity at 2 lb. pressure, that is, the delivery in cubic
feet against 2 lb. pressure, the cylinder blower would cost no more
than, if as much as, the rotary blower.
In this connection it is worth while making a note of the increasing
tendency of lead smelters to provide much more powerful blowers than
were formerly considered necessary, due, no doubt, in large measure to
the recognition of the greater loss of air by leakage backward at the
pressure now worked against. It is considered, for example, that a 42 ×
140 in. furnace to be driven under 40 oz. pressure should be provided
with a No. 10 blower, which size displaces 300 cu. ft. of air per
revolution and is designed to be run at about 100 r.p.m.; its nominal
capacity is, therefore, 30,000 cu. ft. of air per minute; although
its actual delivery against 40 oz. pressure is much less, as pointed
out by Mr. Raht and Mr. Bretherton. The Connersville Blower Company,
of Connersville, Ind., lately supplied the Aguas Calientes plant (now
of the American Smelting and Refining Company) with a rotary blower
of the above capacity, and duplicates of it have been installed at
other smelting works. The force required to drive such a huge blower
is enormous, being something like 400 h.p., which makes it advisable
to provide each blower with a directly connected compound condensing
engine.
In view of the favor with which cylindrical blowing engines for driving
lead blast furnaces are held by many of the leading lead-smelting
engineers, and the likelihood that they will come more and more into
use, it will be interesting to observe whether the lead smelters will
take another step in the tracks of the iron smelters and adopt the
circular form of blast furnace that is employed for the reduction
of iron ore. The limit of size for rectangular furnaces appears to
have been reached in those of 42 × 145 in., or approximately those
dimensions. A furnace of 66 × 160 in., which was built several years
ago at the Globe plant at Denver, proved a failure. H. V. Croll at
that time advocated the building of a circular furnace instead of the
rectangular furnace of those excessive dimensions and considered that
the experience with the latter demonstrated their impracticability. In
the _Engineering and Mining Journal_ of May 28, 1898, he stated that
there was no good reason, however, why a furnace of 300 to 500 tons
daily capacity could not be run successfully, but considered that the
round furnace was the only form permissible. We are unaware whether
Mr. Croll was the first to advocate the use of large circular furnaces
for lead smelting, but at all events there are other experienced
metallurgists who now agree with him, and the time is, perhaps, not far
distant when they may be adopted.
ROTARY BLOWERS VS. BLOWING ENGINES
BY J. PARKE CHANNING
(June 8, 1901)
In the issues of the _Engineering and Mining Journal_ for April
13th and 27th reference was made to the relative efficiency of
piston-blowing engines and rotary blowers of the impeller type, and in
these articles August Raht was quoted as saying that, with an ordinary
rotary blower working against 10 lb. pressure, the loss was 100 per
cent. I have waited some time with the idea that some of the blower
people would call attention to the concealed fallacy in the statement
quoted, but so far have failed to notice any reference to the matter. I
feel quite sure that Mr. Bretherton failed to quote Mr. Raht in full.
The one factor missing in this statement is the speed at which the
blower was run when the loss was 100 per cent.
The accepted method of testing the volumetric efficiency of rotary
blowers is that of “closed discharge.” The discharge opening of the
blower is closed, a pressure gage is connected with the closed delivery
pipe, and the blower is gradually speeded up until the gage registers
the required pressure. The number of revolutions which the blower
makes while holding that pressure, multiplied by the cubic feet per
revolution, will give the total slip of that particular blower at that
particular pressure. Experience has shown that, within the practical
limits of speed at which a blower is run, the slip is a function of
the pressure and has nothing to do with the speed. If, therefore, it
were found that the particular blower referred to by Mr. Raht were
obliged to be revolved at the rate of 30 r.p.m. in order to maintain a
constant pressure of 10 lb. with a closed discharge, and if the blower
were afterward put in practical service, delivering air, and were run
at a speed of 150 r.p.m., it would then follow that its delivery of air
would amount to: 150-30 = 120. Its volumetric efficiency would be 120
÷ 150 = 80 per cent. The above figures must not be relied upon, as I
give them simply by way of illustration.
About a year ago I had the pleasure of examining the tabulated results
of some extensive experiments in this direction, made by one of the
blower companies. I believe they carried their experiments up to 10 lb.
pressure, and I regret that I have not the figures before me, so that
I could give something definite. I do, however, remember that in the
experimental blower, when running at about 150 r.p.m., the volumetric
efficiency at 2 lb. pressure was about 85 per cent., and that at 3 lb.
pressure the volumetric efficiency was about 81 per cent.
It is unnecessary in this connection to call attention to the
horse-power efficiency of rotary blowers. This is a matter entirely
by itself, and there is considerable difference of opinion among
engineers as to the relative horse-power efficiency of rotary blowers
and piston blowers. All agree that there is a certain pressure at which
the efficiency of the blower becomes less than the efficiency of the
blowing engine. This I have heard placed all the way from 2 lb. up to 6
lb.
At the smelting plant of the Tennessee Copper Company we have lately
installed blast-furnace piston-blowing engines; the steam cylinders
are of the Corliss type and are 13 and 24 in. by 42 in.; the blowing
cylinders are two in number, each 57 × 42 in.; the air valves are all
Corliss in type. These blowing engines are designed to operate at a
maximum air pressure of 2½ lb. per square inch.
At the Santa Fe Gold and Copper Mining Company’s smelter we have
recently installed a No. 8 blower directly coupled to a 14 × 32 in.
Corliss engine. This blower has been in use about five months and is
giving very good results against the comparatively low pressure of 12
oz., or ¾ lb.
During the coming summer it is my intention to make careful volumetric
and horse-power tests on these two types of machines under similar
conditions of air pressure, and to publish the results; but in the
meantime I wish to correct the error that a rotary blower of the
impeller type is not a practicable machine at pressure over 5 lb.
BLOWERS AND BLOWING ENGINES FOR LEAD AND COPPER SMELTING
BY HIRAM W. HIXON
(July 20, 1901)
In the _Engineering and Mining Journal_ for July 6th I note the
discussion over the relative merits of blowers and blowing engines for
lead and copper smelting, and wish to state that, irrespective of the
work to be done, the blast pressure will depend entirely on the charge
burden in any kind of blast-furnace work, and that the charge burden
governs the reducing action of the furnace altogether. Along these
lines the iron industry has raised the charge burden up to 100 ft. to
secure the full benefit of the reducing action of the carbon monoxide
on the ore.
In direct opposition to this we have what is known as pyritic smelting,
wherein the charge burden governs the grade of the matte produced to
such an extent that if a charge run with 4 to 6 ft. of burden above the
tuyeres, producing 40 per cent. matte, is changed to a charge burden of
10 or 12 ft., the grade of the matte will decrease from 40 per cent. to
probably less than 20 per cent. I can state this as a fact from recent
experience in operating a blast furnace on heap-roasted ores under the
conditions named, with the result as above stated.
I was exceedingly skeptical about pyritic smelting as advocated by
some of your correspondents, and still continue to be so; but on
making inquiries from some of my co-workers in this line, Mr. Sticht
of Tasmania, and Mr. Nutting of Bingham, Utah, I have arrived at the
following conclusion, to which some may take exception: That pyritic
smelting without fuel, or with less than 5 per cent., with hot
blast, is practically impossible; that smelting raw ore with a low
charge burden to avoid the reducing action of the carbon monoxide,
thereby securing oxidation of the iron and sulphur, is possible and
practicable, under favorable conditions; and that a large portion of
the sulphur is burned off, and the iron, without reducing action,
goes into the slag in combination with silica. These results can be
obtained with cold blast.
A blowing engine would certainly be much out of place for operating
copper-matting furnaces run with the evident intention of oxidizing
sulphur and iron and securing as high a grade of matte as possible,
for the reason that to do this it is necessary to run a low charge
burden, and with a low charge burden a high pressure of blast cannot
be maintained. With a 4 to 6 ft. charge burden the blast pressure at
Victoria Mines at present is 3 oz., produced by a No. 6 Green blower
run at 120 r.p.m.; and a blowing engine, delivering the same amount
of air, would certainly not give more pressure. Under the conditions
which we have, a fan would be more effective than a pressure blower,
and a blowing engine entirely out of the question as far as economy is
concerned.
I installed blowing engines at the East Helena for lead smelting where
the charge burden was 21 ft. and the blast pressure at times went up
as high as 48 oz. Under these conditions the blowing engines gave
satisfaction, but I am of the opinion that the same amount of blast
could have been obtained under that pressure with less horse-power by
the best type of rotary blower. I do not believe that the field of
the blowing engine properly exists below 5 lb., and if this pressure
cannot be obtained by charge-burden conditions, their installation is a
mistake.
I wish to add the very evident fact that varying the grade of the matte
by feeding the furnace at different hights varies the slag composition
as to its silica and iron contents and makes the feeder the real
metallurgist. The reducing action in the furnace is effected almost
entirely by the gases, and when these are allowed to go to waste,
reduction ceases.
BLOWING ENGINES AND ROTARY BLOWERS—HOT BLAST FOR PYRITIC SMELTING
BY S. E. BRETHERTON
(August 24, 1901)
I have just read in the _Engineering and Mining Journal_ of July 20th
an interesting letter written by Hiram W. Hixon in regard to blowing
engines versus the rotary blowers, and also the use of cold blast for
pyritic smelting.
The controversy, which I unintentionally started in my letter in
the _Engineering and Mining Journal_ of April 13th last, about the
advantages of using either blowers or blowing engines for blast
furnaces, does not particularly interest me, with the exception that I
have about decided, in my own mind, to use blowing engines where there
is much back pressure, and the ordinary up-to-date blower for pyritic
or matte smelting where much back pressure should not exist. I fully
appreciate the fact that so-called pyritic smelting can be done to a
limited extent, even with cold blast. Theoretically, enough oxygen
can be sent into the blast furnace, contained in the cold blast, to
oxidize both the fuel and the sulphur in an ordinary sulphide charge,
but I have not yet learned where a high concentration is being made
with unroasted ore and cold blast. I experimented on these lines at
different times for three years, during 1896, 1897, and 1898, making
a fair concentration with refractory ores, most of which had been
roasted. I was myself interested in the profits and as anxious as any
one for economy. We tried, for fuel in the blast furnace, coke alone,
coke and lignite coal, lignite coal alone, lignite coal and dry wood,
coal and green wood, and then coke and green wood, all under different
hights of ore burden in the furnace.
A description of these experiments would, no doubt, be tiresome to your
readers, but I wish to state that the furnace was frozen up several
times on account of using too little fuel, when the cold blast would
gradually drive nearly all the heat to the top of the furnace, the
crucible and between the tuyeres becoming so badly crusted that the
furnace had to be cleaned out and blown in again, unless I was called
in time to save it by changing the charge and increasing the fuel. We
were making high-grade matte under contract, high concentration and
small matte fall, which would, no doubt, aggravate matters.
After the introduction of hot blast, heated up to between 200 and 300
deg. F., we made the same grade of matte from the same character of
ore, with the exception that we then smelted without roasting, and
reduced the percentage of fuel consumption, increased the capacity of
the furnace, and almost entirely obviated the trouble of cold crucibles
and hot tops. I write the above facts, as they speak for themselves.
I nearly agree with Mr. Hixon, and do not think it practical to smelt
with much less than 5 per cent. coke continuously; but there is a
great saving between the amount of coke used with a moderately heated
blast and cold blast. Regardless of either hot or cold blast, however,
the fuel consumption depends very much on the character of the ore
to be smelted, the amount of matte-fall and grade of matte made. It
is not always advisable or necessary to use hot blast for a matting
furnace; that is, where the supply of sulphur is limited. It may then
be necessary to use as much fuel in the blast furnace to prevent the
sulphur from oxidizing as will be sufficient to furnish the heat for
smelting. Such conditions existed at Silver City, N. M. , at times,
after our surplus supply of iron and zinc sulphide concentrates was
used. I understand that they are now short of sulphur there, on account
of getting a surplus amount of oxidized copper ore, and are only
utilizing what little heat the slag gives them, without the addition of
any fuel on top of the forehearth.
Before closing this, which I intended to have been brief, I wish to
call your attention to a little experience we had with alumina in the
matting furnace at Silverton, Col., where I was acting as consulting
metallurgist. The ore we had to smelt contained, on an average, about
20 per cent. Al₂O₃, 30 per cent. SiO₂, with 18 per cent. Fe in
the form of an iron pyrite, and no other iron was available except some
iron sulphide concentrates containing a small percentage of zinc and
lead.
The question naturally arose, could we oxidize and force sufficient
of the iron into the slag, and where should we class the alumina, as
a base or an acid? My experience in lead smelting led me to believe
that Al₂O₃ could only be classed as an acid in the ordinary
lead furnace, and that it would be useless to class it otherwise in
a shallow matting furnace; and E. W. Walter, the superintendent and
metallurgist in charge, agreed with me.
We then decided to make a bisilicate slag, classing the alumina as
silica, and we obtained fairly satisfactory results. The slag made
was very clean, but treacherous, which was attributed to two reasons:
First, that it required more heat to keep the alumina slag liquid
enough to flow than it does a nearly straight silica slag; and, second,
that we were running so close to the formula of a bisilicate and
aluminate slag (about 31½ per cent. SiO₂, 27 per cent. Fe, 20 per
cent. CaO, and 18 per cent. Al₂O₃, or 49½ per cent. acid) that a
few charges thrown into the furnace containing more silica or alumina
than usual would thicken the slag so that it would then require some
extra coke and flux to save the furnace. At times the combined SiO₂
and Al₂O₃ did reach 55 and 56 per cent. in the slag, which did
not freeze up the furnace, but caused us trouble.
PART IX
LEAD REFINING
THE REFINING OF LEAD BULLION[49]
BY F. L. PIDDINGTON
(October 3, 1903)
In presenting this account of the Parkes process of desilverizing and
refining lead bullion no originality is claimed, but I hope that a
description of the process as carried out at the works of the Smelting
Company of Australia may be of service.
_Introductory._—The Parkes process may be conveniently summarized as
follows:
1. Softening of the base bullion to remove copper, antimony, etc.
2. Removal of precious metals from the softened bullion by means of
zinc.
3. Refining the desilverized lead.
4. Liquation of gold and silver crusts obtained from operation No. 2.
5. Retorting the liquated alloy to drive off zinc.
6. Concentrating and refining bullion from No. 5.
_Softening._—This is done in reverberatory furnaces. In large works two
furnaces are used, copper, antimony, and arsenic being removed in the
first and antimony in the second. The size of the furnaces is naturally
governed by the quantity to be treated. In these works (refining about
200 tons weekly) a double set of 15-ton furnaces were at work. The
sides and ends of these furnaces are protected by a jacket with a 2-in.
water space, the jacket extending some 3 in. above the charge level and
6 in. to 9 in. below it. The furnace is built into a wrought-iron pan,
and if the brickwork is well laid into the pan there need be no fear of
lead breaking through below the jacket.
The bars of bullion (containing, as a rule, 2 to 3 per cent. of
impurities) are placed in the furnace carefully, to avoid injuring the
hearth, and melted down slowly. The copper dross separates out and
floats on top of the charge, which is stirred frequently to expose
fresh surfaces. If the furnace is overheated some dross is melted into
the lead again and will not separate out until the charge is cooled
back. However carefully the work is done some copper remains with the
lead, and its effects are to be seen in the later stages. The dross is
skimmed into a slag pot with a hole bored in it some 4 in. from the
bottom; any lead drained from the pot is returned to the charge. The
copper dross is either sent back to the blast furnace direct or may
be first liquated. By the latter method some 30 per cent. of the lead
contents of the dross is recovered in the refinery.
Base bullion made at a customer’s smelter will often vary greatly in
composition, and it is, therefore, difficult to give any hard and fast
figures as to percentage of metals in the dross. As a rule our dross
showed 65 to 70 per cent. lead, copper 2 to 9 per cent. (average 4 per
cent.), gold and silver values varying with the grade of the original
bullion, though it was difficult to detect any definite relation
between bullion and dross. It was, however, noticed that gold and
silver values increased with the percentage of copper.
Immediately the copper dross is skimmed off the heat is raised
considerably, and very soon a tin (and arsenic, if present) skimming
appears. It is quite “dry” and may be removed in an hour or so. It is a
very small skimming, and the tin, not being worth saving, is put with
the copper dross.
The temperature is now raised again and antimony soon shows in black,
boiling, oily drops, gathering in time into a sheet covering the
surface of the lead. When the skimming is about ½-inch thick, slaked
lime, ashes, or fine coal is thrown on and stirred in. The dross soon
thickens up and may be skimmed off easily. This operation is repeated
until all antimony is eliminated. Constant stirring of the charge is
necessary. The addition of litharge greatly facilitates the removal
of antimony; either steam or air may be blown on the surface of the
metal to hasten oxidation, though they have anything but a beneficial
effect on the furnace lining. From time to time samples of the dross
are taken in a small ladle, and after setting hard the sample is broken
in two. A black vitreous appearance indicates plenty of antimony yet
in the charge. Later samples will look less black, until finally a few
yellowish streaks are seen, being the first appearance of litharge.
When all antimony is out the fracture of a sample should be quite
yellow and the grain of the litharge long, a short grain indicating
impurities still present, in which case another skimming is necessary.
The analysis of a representative sample of antimony dross was as
follows:
PbO = 78.11 per cent.
Sb₂O₄ = 8.75 ” ”
As₂O₃ = 2.18 ” ”
CuO = 0.36 ” ”
CaO = 1.10 per cent.
Fe₂O₃ = 0.42 ” ”
Al₂O₃ = 0.87 ” ”
Insol. = 4.10 ” ”
Antimony dross is usually kept separate and worked up from time to
time, yielding hard antimonial lead, used for type metal, Britannia
metal, etc.
_Desilverization._—The softening being completed the charge is tapped
and run to a kettle or pan of cast iron or steel, holding, when
conveniently full, some 12 or 13 tons. The lead falling into the
kettle forms a considerable amount of dross, which is skimmed off and
returned to the softening furnace. By cooling down the charge until
it nearly “freezes” an additional copper skimming is obtained, which
also is returned to the softener. The kettle is now heated up to the
melting point of zinc, and the zinc charge, determined by the gold
and silver contents of the kettle, is added and melted. The charge
is stirred, either by hand or steam, for about an hour, after which
the kettle is allowed to cool down for some three hours and the first
zinc crust taken off. When the charge is skimmed clean a sample of the
bullion is taken for assay, and while this is being done the kettle is
heated again for the second zinc charge, which is worked in the same
way as the first; sometimes a third addition of zinc is necessary. The
resulting crusts are kept separate, the second and third being added
to the next charge as “returns,” allowing 3 lb. of zinc in returns as
equal to 1 lb. of fresh zinc. An alternative method is to take out gold
and silver in separate crusts, in which case the quantity of zinc first
added is calculated on the gold contents of the kettle only. The method
of working is the same, though subsequent treatment may differ in that
the gold crusts are cupeled direct.
As to the quantity of zinc required:
1. Extracting the gold with as little silver as possible, the following
figures were obtained:
Total Gold— Au.
In kettle 300 oz. │ 1 lb. zinc takes out 1.00 oz.
” ” 200 ” │ ” ” ” ” 1.00 ”
” ” 150 ” │ ” ” ” ” 0.79 ”
” ” 100 ” │ ” ” ” ” 0.59 ”
” ” 60 ” │ ” ” ” ” 0.45 ”
2. Silver zinking gave the following general results with 11-ton
charges:
Total Silver—
In kettle 1,450 oz. │ 1 lb. zinc takes out 5.6 oz.
” ” 1,200 ” │ ” ” ” ” 4.1 ”
” ” 930 ” │ ” ” ” ” 3.8 ”
” ” 755 ” │ ” ” ” ” 3.5 ”
” ” 616 ” │ ” ” ” ” 3.4 ”
” ” 460 ” │ ” ” ” ” 2.6 ”
3. Extracting gold and silver together:
───────────────────────────┬──────────────────────
TOTAL CONTENTS OF KETTLE │ 1 LB. ZINC TAKES OUT
AU. OZ. │ AG. OZ. │ AU. OZ. │ AG. OZ.
────────────┼──────────────┼─────────────┼────────
494 │ 3,110 │ 0.59 │ 3.60
443 │ 1,883 │ 0.64 │ 2.80
330 │ 2,417 │ 0.45 │ 3.34
204 │ 1,638 │ 0.36 │ 2.86
143 │ 1,330 │ 0.28 │ 2.65
123 │ 1,320 │ 0.23 │ 2.54
────────────┴──────────────┴─────────────┴────────
It will be noticed that in each case the richer the bullion the greater
the extractive power of zinc. Experiments made on charges of rich
bullion showed that the large amount of zinc called for by the table in
use was unnecessary, and 250 lb. was fixed on as the first addition of
zinc. On this basis an average of 237 charges gave results as follows:
───────────────────┬───────────┬──────────────────────
TOTAL CONTENTS │ ZINC USED │ 1 LB. ZINC TAKES OUT
AU. OZ. │ AG. OZ. │ LBS. │ AU. OZ. │ AG. OZ.
────────┼──────────┼───────────┼───────────┼──────────
520 │ 1,186 │ 507.5 │ 1.27 │ 2.91
────────┴──────────┴───────────┴───────────┴──────────
The zinc used was that necessary to clean the kettle, added as follows:
1st, 250 lb.; 2d (average), 127 lb.; 3d (average), 57 lb. In 112 cases
no third addition was required. From these figures it appears that in
the earlier work the zinc was by no means saturated.
_Refining the Lead._—Gold and silver being removed, the lead is
siphoned off into the refining kettle and the fire made up. In about
four hours the lead will be red hot, and when hot enough to burn zinc,
dry steam, delivered by a ¾ in. pipe reaching nearly to the bottom of
the kettle, is turned on. The charge is stirred from time to time and
wood is fed on the top to assist dezinking and prevent the formation
of too much litharge. In three to four hours the lead will be soft and
practically free from zinc. When test strips show the lead to be quite
soft and clean, the kettle is cooled down and the scum of lead and
zinc oxides skimmed off. In an hour or so the lead will be cool enough
for molding; the bar should have a yellow luster on the face when set;
if the lead is too cold it will be white, if too hot a deep blue. The
refining kettles are subjected to severe strain during the steaming
process, and hence their life is uncertain—an average would be about 60
charges; the zinking kettles, on the other hand, last very much longer.
Good steel kettles (if they can be obtained) are preferable to cast
iron.
_Treatment of Zinc Crusts._—Having disposed of the lead, let us
return now to the zinc crusts. These are first liquated in a small
reverberatory furnace, the hearth of which is formed of a cast-iron
plate (the edges of the long sides being turned up some 4 in.) laid on
brasque filling, with a fall from bridge to flue of ¾ in. per foot and
also sloping from sides to center. The operation is conducted at a low
temperature and the charge is turned over at intervals, the liquated
lead running out into a small separately fired kettle. This lead rarely
contains more than a few ounces of silver per ton; it is baled into
bars, and returned to the zinking kettles or worked up in a separate
charge. In two to three hours the crust is as “dry” as it is advisable
to make it, and the liquated alloy is raked out over a slanting
perforated plate to break it up and goes to the retort bin.
_Retorting the Alloy._—This is carried on in Faber du Faur tilting
furnaces—simply a cast-iron box swinging on trunnions and lined with
firebrick. Battersea retorts (class 409) holding 560 lb. each are
used; their average life is about 30 charges. The retorts are charged
hot, a small shovel of coal being added with the alloy. The condenser
is now put in place and luted on; it is made of ⅛ in. iron bent to
form a cylinder 12 in. in diameter, open at one end; it is lined with
a mixture of lime, clay and cement. It has three holes, one on the
upper side close to the furnace and through which a rod can be passed
into the retort, a vent-hole on the upper side away from the furnace,
and a tap-hole on the bottom for condensed zinc. In an hour or so the
flame from the vent-hole should be green, showing that distillation has
begun. When condensation ceases (shown by the flame) the condenser is
removed and the bullion skimmed and poured into bars for the cupel. The
products of retorting are bullion, zinc, zinc powder and dross. Bullion
goes to the cupel, zinc is used again in the desilverizing kettles,
powder is sieved to take out scraps of zinc and returned to the blast
furnace, or it may be, and sometimes is, used as a precipitating
agent in cyanide works; dross is either sweated down in a cupel with
lead and litharge, together with outside material such as zinc gold
slimes from cyanide works, jeweler’s sweep, mint sweep, etc., or in
the softening furnace after the antimony has been taken off. In either
case the resulting slag goes back to the blast furnace. The total
weight of alloy treated is approximately 7 per cent. of the original
base bullion. The zinc recovered is about 60 per cent. of that used
in desilverizing. The most important source of temporary loss is the
retort dross (consisting of lead-zinc-copper alloy with carbon, silica
and other impurities), and it is here that the necessity of removing
copper in the softening process is seen, since any copper comes out
with the zinc crusts and goes on to the retorts, where it enters the
dross, carrying gold and silver with it. If much copper is present the
dross may contain more gold and silver than the retort bullion itself.
In this connection I remember an occasion on which some retort dross
yielded gold and silver to the extent of over 800 and 3000 oz. per ton
respectively.
_Cupellation._—Retort bullion is first concentrated (together with
bullion resulting from dross treatment) to 50 to 60 per cent. gold and
silver in a water-jacketed cupel. The side lining is protected by an
inch water-pipe imbedded in the lining at the litharge level or by a
water-jacket, the inner face of which is of copper; the cupel has also
a water-jacketed breast so that the front is not cut down. The cupel
lining may be composed of limestone, cement, fire-clay and magnesite
in various proportions, but a simple lining of sand and cement was
found quite satisfactory. When the bullion is concentrated up to 50 to
60 per cent. gold and silver, it is baled out and transferred to the
finishing cupel, where it is run up to about 0.995 fine; it is then
ready either for the melting-pot or parting plant. The refining test,
by the way, is not water-cooled.
Re-melting is done in 200-oz. plumbago crucibles and presents no
special features. In the case of doré bullion low in gold, “sprouting”
of the silver is guarded against by placing a piece of wood or charcoal
on the surface of the metal before pouring, and any slag is kept back.
The quantity of slag formed is, of course, very small, so that the bars
do not require much cleaning.
The parting plant was not in operation in my time, and I am therefore
unable to go into details. The process arranged for was briefly as
follows: Solution of the doré bullion in H₂SO₄; crystallization
of silver as monosulphate by dilution and cooling; decomposition of
silver sulphate by ferrous sulphate solution giving metallic silver and
ferric sulphate, which is reduced to the ferrous salt by contact with
scrap iron. The gold and silver are washed thoroughly with hot water
and cast into bars.
In conclusion, some variations in practice may be noted. The use of
two furnaces in the softening process has already been mentioned; by
this means the drossing and softening are more perfect and subsequent
operations thereby facilitated; further, the furnaces, being kept at
a more equable temperature, are less subject to wear and tear. Zinc
crusts are sometimes skimmed direct into an alloy press in which
the excess of lead is squeezed out while still molten; liquation is
then unnecessary. Refining of the lead may be effected by a simple
scorification in a reverberatory, the soft lead being run into a kettle
from which it is molded into market bars.
THE ELECTROLYTIC REFINING OF BASE LEAD BULLION
BY TITUS ULKE
(October 11, 1902)
Important changes in lead-refining practice are bound to follow, in my
opinion, the late demonstration on a large scale of the low working
cost and high efficiency of Betts’ electrolytic process of refining
lead bullion. It was my good fortune recently to see this highly
interesting process in operation at Trail, British Columbia, through
the kindness of the inventor, A. G. Betts, and Messrs. Labarthe and
Aldridge, of the Trail works.
A plant of about 10 tons daily capacity, which probably cost about
$25,000, although it could be duplicated for perhaps $15,000 at the
present time, was installed near the Trail smelting works. It has been
in operation for about ten months, I am informed, with signal success,
and the erection of a larger plant, of approximately 30 tons capacity
and provided with improved handling facilities, is now completed.
The depositing-room contains 20 tanks, built of wood, lined with tar,
and approximately of the size of copper-refining tanks. Underneath the
tank-room floor is a basement permitting inspection of the tank bottoms
for possible leakage and removal of the solution and slime. A suction
pump is employed in lifting the electrolyte from the receiving tank and
circulating the solution. In nearly every respect the arrangement of
the plant and its equipment is strikingly like that of a modern copper
refinery.
The great success of the process is primarily based upon Betts’
discovery of the easy solubility of lead in an acid solution of lead
fluosilicate, which possesses both stability under electrolysis and
high conductivity, and from which exceptionally pure lead may be
deposited with impure anodes at a very low cost. With such a solution,
there is no polarization from formation of lead peroxide on the anode,
no evaporation of constituents except water, and no danger in its
handling. It is cheaply obtained by diluting hydrofluoric acid of
35 per cent. HF, which is quoted in New York at 3c. per pound, with
an equal volume of water and saturating it with pulverized quartz
according to the equation:
SiO₂ + 6HF = HSiF₆ + 2H₂O.
According to Mr. Betts, an acid of 20 to 22 per cent. will come
to about $1 per cu. ft., or to $1.25 when the solution has been
standardized with 6 lb. of lead. One per cent. of lead will neutralize
0.7 per cent. H₂SiF₆. The electrolyte employed at the time of my
inspection of the works contained, I believe, 8 per cent. lead and 11
per cent. excess of fluosilicic acid.
The anodes consist of the lead bullion to be refined, cast into plates
about 2 in. thick and approximately of the same size as ordinary
two-lugged copper anodes. Before being placed in position in the tanks,
they are straightened by hammering over a mold and their lugs squared.
No anode sacks are employed as in the old Keith process.
The cathode sheets which receive the regular lead deposits are thin
lead plates obtained by electrodeposition upon and stripping from
special cathodes of sheet steel. The latter are prepared for use by
cleaning, flashing with copper, lightly lead-plating in the tanks, and
greasing with a benzine solution of paraffin, dried on, from which the
deposited lead is easily stripped.
The anodes and cathodes are separated by a space of 1½ to 2 in. in the
tank and are electrically connected in multiple, the tanks being in
series circuit. The fall in potential between tanks is only about 0.2
of a volt, which remarkably low voltage is due to the high conducting
power of the electrolyte and to some extent to the system of contacts
used. These contacts are small wells of mercury in the bus-bars, large
enough to accommodate copper pins soldered to the iron cathodes or
clamped to the anodes. Only a small amount of mercury is required.
Current strengths of from 10 to 25 amperes per sq. ft. have been used,
but at Trail 14 amperes have given the most satisfactory results as
regards economy of working and the physical and chemical properties of
the refined metal produced.
A current of 1 ampere deposits 3.88 grams of lead per hour, or
transports 3¼ times as much lead, in this case, as copper with an
ordinary copper-refining solution. A little over 1000 kg., or 2240
lb., requires about 260,000 ampere hours. At 10 amperes per sq. ft. the
cathode (or anode) area should be about 1080 sq. ft. per ton of daily
output. Taking a layer of electrolyte 1.5 in. thick, 135 cu. ft. will
be found to be the amount between the electrodes, and 175 cu. ft. may
be taken as the total quantity of solution necessary, according to Mr.
Betts’ estimate. The inventor states that he has worked continuously
and successfully with a drop of potential of only 0.175 volt per tank,
and that therefore 0.25 volt should be an ample allowance in regular
refining. Quoting Mr. Betts; “260,000 ampere hours at 0.25 volt works
out to 87 electrical h.p. hours of 100 h.p. hours at the engine shaft,
in round numbers. Estimating that 1 h.p. hour requires the burning of
1.5 lb. of coal, and allowing say 60 lb. for casting the anodes and
refined lead, each ton of lead refined requires the burning of 210 lb.
of fuel.” With coal at $6 per ton the total amount of fuel consumed,
therefore, should not cost over 60c., which is far below the cost of
fire-refining base lead bullion, as we know.
In the Betts electrolytic process, practically all the impurities
in the base bullion remain as a more or less adherent coating on
the anode, and only the zinc, iron, cobalt and nickel present go
into solution. The anode residue consists practically of all the
copper, antimony, bismuth, arsenic, silver and gold contained in the
bullion, and very nearly 10 per cent. of its weight in lead. Having
the analysis of any bullion, it is easy to calculate with these data
the composition of the anode residue and the rate of pollution of the
electrolyte. Allowing 175 cu. ft. of electrolyte per ton of daily
output, it will be found that in the course of a year these impurities
will have accumulated to the extent of a very few per cent. Estimating
that the electrolyte will have to be purified once a year, the amount
to be purified daily is less than 1 cu. ft. for each ton of output.
The amount of lead not immediately recovered in pure form is about
0.3 per cent., most of which is finally recovered. As compared with
the ordinary fire-refined lead, the electrolytically refined lead is
much purer and contains only mere traces of bismuth, when bismuthy
base bullion is treated. Furthermore, the present loss of silver in
fire refining, amounting, it is claimed, to about 1½ per cent. of the
silver present, and covered by the ordinary loss in assay, is to a
large extent avoided, as the silver in the electrolytic process is
concentrated in the anode residue with a very small loss, and the loss
of silver in refining the slimes is much less than in treating the
zinc crusts and refining the silver residue after distillation. The
silver slimes obtained at Trail, averaging about 8000 oz. of gold and
silver per ton, are now treated at the Seattle Smelting and Refining
Works. There the slimes are boiled with concentrated sulphuric acid and
steam, allowing free access of air, which removes the greater part of
the copper. The washed residue is then dried in pans over steam coils,
and melted down in a magnesia brick-lined reverberatory, provided
with blast tuyeres, and refined. In this reverberatory furnace the
remainder of the copper left in the slimes after boiling is removed by
the addition of niter as a flux, and the antimony with soda. The doré
bars finally obtained are parted in the usual way with sulphuric acid,
making silver 0.999 fine and gold bars at least 0.992 fine.
Mr. Betts treated 2000 grams of bullion, analyzing 98.76 per cent. Pb,
0.50 Ag, 0.31 Cu, and 0.43 Sb with a current of 25 amp. per square
foot in an experimental way, and obtained products of the following
composition:
Refined Lead: 99.9971 per cent. Pb, 0.0003 Ag, 0.0007 Cu, and 0.0019 Sb.
Anode Residue: 9.0 per cent. Pb, 36.4 Ag, 25.1 Cu, and 2.95 Sb.
Four hundred and fifty pounds of bullion from the Compania Metalurgica
Mexicana, analyzing 0.75 per cent. Cu, 1.22 Bi, 0.94 As, 0.68 Sb, and
assaying 358.9 oz. Ag and 1.71 oz. Au per ton, were refined with a
current of 10 amp. per square foot, and gave a refined lead of the
following analysis: 0.00027 per cent. Cu, 0.0037 Bi, 0.0025 As, 0 Sb,
0.0010 Ag, 0.0022 Fe, 0.0018 Zn and Pb (by difference) 99.9861 per cent.
Although the present method for recovering the precious metals and
by-products from the anode residue leaves much room for improvement,
the use of the Betts process may be recommended to our lead refiners,
because it is a more economical and efficient method than the
fire-refining process now in common use. I will state my belief, in
conclusion, that the present development of electrolytic lead refining
signalizes as great an advance over zinc desilverization and the fire
methods of refining lead as electrolytic copper refining does over the
old Welsh method of refining that metal.
ELECTROLYTIC LEAD-REFINING[50]
BY ANSON G. BETTS
A solution of lead fluosilicate, containing an excess of fluosilicic
acid, has been found to work very satisfactorily as an electrolyte
for refining lead. It conducts the current well, is easily handled
and stored, non-volatile and stable under electrolysis, may be made
to contain a considerable amount of dissolved lead, and is easily
prepared from inexpensive materials. It possesses, however, in common
with other lead electrolytes, the defect of yielding a deposit of lead
lacking in solidity, which grows in crystalline branches toward the
anodes, causing short circuits. But if a reducing action (practically
accomplished by the addition of gelatine or glue) be given to the
solution, a perfectly solid and dense deposit is obtained, having very
nearly the same structure as electrolytically deposited copper, and a
specific gravity of about 11.36, which is that of cast lead.
Lead fluosilicate may be crystallized in very soluble brilliant
crystals, resembling those of lead nitrate and containing
four molecules of water of crystallization, with the formula
PbSiF₆,4H₂O. This salt dissolves at 15 deg. C. in 28 per cent. of
its weight of water, making a syrupy solution of 2.38 sp. gr. Heated
to 60 deg. C., it melts in its water of crystallization. A neutral
solution of lead fluosilicate is partially decomposed on heating, with
the formation of a basic insoluble salt and free fluosilicic acid,
which keeps the rest of the salt in solution. This decomposition ends
when the solution contains perhaps 2 per cent. of free acid; and the
solution may then be evaporated without further decomposition. The
solutions desired for refining are not liable to this decomposition,
since they contain much more than 2 per cent. of free acid. The
electrical conductivity depends mainly on the acidity of the solution.
My first experiments were carried out without the addition of gelatine
to the fluosilicate solution. The lead deposit consisted of more or
less separate crystals that grew toward the anode, and, finally, caused
short circuits. The cathodes, which were sheet-iron plates, lead-plated
and paraffined, had to be removed periodically from the tanks and
passed through rolls, to pack down the lead. When gelatine has been
added in small quantities, the density of the lead is greater than can
be produced by rolling the crystalline deposit, unless great pressure
is used.
The Canadian Smelting Works, Trail, B. C. , have installed a refinery,
making use of this process. There are 28 refining-tanks, each 86 in.
long, 30 in. wide and 42 in. deep, and each receiving 22 anodes of
lead bullion with an area of 26 by 33 in. exposed to the electrolyte
on each side, and 23 cathodes of sheet lead, about 1/16 in. thick,
prepared by deposition on lead-plated and paraffined iron cathodes. The
cathodes are suspended from 0.5 by 1 in. copper bars, resting crosswise
on the sides of the tanks. The experiment has been thoroughly tried of
using iron sheets to receive a deposit thicker than 1/16 in.; that is,
suitable for direct melting without the necessity of increasing its
weight by further deposition as an independent cathode; but the iron
sheets are expensive, and are slowly pitted by the action of the acid
solution; and the lead deposits thus obtained are much less smooth and
pure than those on lead sheets.
The smoothness and the purity of the deposited lead are proportional.
Most of the impurity seems to be introduced mechanically through the
attachment of floating particles of slime to irregularities on the
cathodes. The effect of roughness is cumulative; it is often observed
that particles of slime attract an undue amount of current, resulting
in the lumps seen in the cathodes. Samples taken at the same time
showed from 1 to 2.5 oz. silver per ton in rough pieces from the iron
cathodes, 0.25 oz. as an average for the lead-sheet cathodes, and only
0.04 oz. in samples selected for their smoothness. The variation in
the amount of silver (which is determined frequently) in the samples
of refined lead is attributed not to the greater or less turbidity of
the electrolyte at different times, but to the employment of new men in
the refinery, who require some experience before they remove cathodes
without detaching some slime from the neighboring anodes.
Each tank is capable of yielding, with a current of 4000 amperes,
750 lb. of refined lead per day. The voltage required to pass this
current was higher than expected, as explained below; and for this
reason, and also because the losses of solution were very heavy until
proper apparatus was put in to wash thoroughly the large volume of
slime produced (resulting in a weakened electrolyte), the current used
has probably averaged about 3000 amperes. The short circuits were
also troublesome, though this difficulty has been greatly reduced by
frequent inspection and careful placing of the electrodes. At one time,
the solution in use had the following composition in grams per 100
c.c.: Pb, 6.07; Sb, 0.0192; Fe, 0.2490; SiF₆, 6.93, and As, a trace.
The current passing was 2800 amperes, with an average of about 0.44
volts per tank, including bus-bars and contacts. It is not known what
was the loss of efficiency on that date, due to short circuits; and
it is, therefore, impossible to say what resistance this electrolyte
constituted.
Hydrofluoric acid of 35 per cent., used as a starting material for the
preparation of the electrolyte, is run by gravity through a series of
tanks for conversion into lead fluosilicate. In the top tank is a layer
of quartz 2 ft. thick, in passing through which the hydrofluoric acid
dissolves silica, forming fluosilicic acid. White lead (lead carbonate)
in the required quantity is added in the next tank, where it dissolves
readily and completely with effervescence. All sulphuric acid and any
hydrofluoric acid that may not have reacted with silica settle out
in combination with lead as lead sulphate and lead fluoride. Lead
fluosilicate is one of the most soluble of salts; so there is never
any danger of its crystallizing out at any degree of concentration
possible under this method. The lead solution is then filtered and run
by gravity into the refining-tanks.
The solution originally used at Trail contained about 6 per cent. Pb
and 15 per cent. SiF₆.
The electrical resistance in the tanks was found to be greater than
had been calculated for the same solution, plus an allowance for
loss of voltage in the contacts and conductors. This is partly, at
least, due to the resistance to free motion of the electrolyte, in
the neighborhood of the anode, offered by a layer of slime which may
be anything up to ½ in. thick. During electrolysis, the SiF₆ ions
travel toward the anodes, and there combine with lead. The lead and
hydrogen travel in the opposite direction and out of the slime; but
there are comparatively few lead ions present, so that the solution
in the neighborhood of the anodes must increase in concentration and
tend to become neutral. This greater concentration causes an e.m.f. of
polarization to act against the e.m.f. of the dynamo. This amounted
to about 0.02 volt for each tank. The greater effect comes from the
greater resistance of the neutral solution with which the slime is
saturated. There is, consequently, an advantage in working with rather
thin anodes, when the bullion is impure enough to leave slime sticking
to the plates. A compensating advantage is found in the increased ease
of removing the slime with the anodes, and wiping it off the scrap in
special tanks, instead of emptying the tanks and cleaning out, as is
done in copper refineries.
It is very necessary to have adequate apparatus for washing solution
out of the slime. The filter first used consisted of a supported
filtering cloth with suction underneath. It was very difficult to
get this to do satisfactory work by reason of the large amount of
fluosilicate to be washed out with only a limited amount of water.
At the present time the slime is first stirred up with the ordinary
electrolyte several times, and allowed to settle, before starting to
wash with water at all. The Trail plant produces daily 8 or 10 cu. ft.
of anode residue, of which over 90 per cent. by volume is solution.
The evaporation from the total tank surface of something like 400 sq.
ft. is only about 15 cu. ft. daily; so that only a limited amount of
wash-water is to be used—namely, enough to replace the evaporated
water, plus the volume of the slime taken out.
The tanks are made of 2 in. cedar, bolted together and thoroughly
painted with rubber paint. Any leakage is caught underneath on sloping
boards. Solution is circulated from one tank to another by gravity, and
is pumped from the lowest to the highest by means of a wooden pump. The
22 anodes in each tank together weigh about 3 tons, and dissolve in
from 8 to 10 days, two sets of cathodes usually being used with each
set of anodes. While 300 lb. cathodes can be made, the short-circuiting
gets so troublesome with the spacing used that the loss of capacity is
more disadvantageous than the extra work of putting in and taking out
more plates. The lead sheets used for cathodes are made by depositing
about 1/16 in. metal on paraffined steel sheets in four of the tanks,
which are different from the others only in being a little deeper.
The anodes may contain any or all of the elements, gold, silver,
copper, tin, antimony, arsenic, bismuth, cadmium, zinc, iron, nickel,
cobalt and sulphur. It would be expected that gold, silver, copper,
antimony, arsenic and bismuth, being more electronegative than lead,
would remain in the slime in the metallic state, with, perhaps, tin,
while iron, zinc, nickel and cobalt would dissolve. It appears that tin
stands in the same relation to lead that nickel does to iron, that is,
they have about the same electromotive forces of solution, with the
consequence that they can behave as one metal and dissolve and deposit
together. Iron, contrary to expectation, dissolves only slightly, while
the slime will carry about 1 per cent. of it. It appears from this that
the iron exists in the lead in the form of matte. Arsenic, antimony,
bismuth and copper have electromotive forces of solution more than 0.3
volt below that of lead. As there is no chance that any particle of
one of these impurities will have an electric potential of 0.3 volt
above that of the lead with which it is in metallic contact, there is
no chance that they will be dissolved by the action of the current. The
same is even more certainly true of silver and gold. The behavior of
bismuth is interesting and satisfactory. It is as completely removed by
this process of refining as antimony is. No other process of refining
lead will remove this objectionable impurity so completely. Tin has
been found in the refined lead to the extent of 0.02 to 0.03 per cent.
This we had no difficulty in removing from the lead by poling before
casting. There is always a certain amount of dross formed in melting
down the cathodes; and the lead oxide of this reacts with the tin in
the lead at a comparatively low temperature.
The extra amount of dross formed in poling is small, and amounts to
less than 1 per cent. of the lead. The dross carries more antimony and
arsenic than the lead, as well as all the tin. The total amount of
dross formed is about 4 per cent. Table I shows its composition.
The electrolyte takes up no impurities, except, possibly, a small part
of the iron and zinc. Estimating that the anodes contain 0.01 per
cent. of zinc and soluble iron, and that there are 150 cu. ft. of the
solution in the refinery for every ton of lead turned out daily, in
one year the 150 cu. ft. will have taken up 93 lb. of iron and zinc,
or about 1 per cent. These impurities can accumulate to a much greater
extent than this before their presence will become objectionable. It
is possible to purify the electrolyte in several ways. For example,
the lead can be removed by precipitation with sulphuric acid, and
the fluosilicic acid precipitated with salt as sodium fluosilicate.
By distillation with sulphuric acid the fluosilicic acid could be
recovered, this process, theoretically, requiring but one-third as much
sulphuric acid as the decomposition of fluorspar, in which the fluorine
was originally contained.
The only danger of lead-poisoning to which the workmen are exposed
occurs in melting the lead and casting it. In this respect the
electrolytic process presents a distinct sanitary advance.
For the treatment of slime, the only method in general use consists in
suspending the slime in a solution capable of dissolving the impurities
and supplying, by a jet of steam and air forced into the solution, the
air necessary for its reaction with, and solution of, such an inactive
metal as copper. After the impurities have been mostly dissolved, the
slime is filtered off, dried and melted, under such fluxes as soda, to
a doré bullion.
The amount of power required is calculated thus: Five amperes in 24
hours make 1 lb. of lead per tank. One ton of lead equals 10,000
ampere-days, and at 0.35 volts per tank, 3500 watt-days, or 4.7
electric h.p. days. Allowing 10 per cent. loss of efficiency in the
tanks (we always get less lead than the current which is passing would
indicate), and of 8 per cent. loss in the generator, increases this to
about 5.6 h.p. days, and a further allowance for the electric lights
and other applications gives from 7 to 8 h.p. days as about the amount
per ton of lead. At $30 per year, this item of cost is something like
65c. per ton of lead. So this is an electro-chemical process not
especially favored by water-power.
The cost of labor is not greater than in the zinc-desilverization
process. A comparison between this process and the Parkes process, on
the assumption that the costs for labor, interest and general expenses
are about equal, shows that about $1 worth of zinc and a considerable
amount of coal and coke have been done away with, at the expense
of power, equal to about 175 h.p. hours, of the average value of
perhaps 65c., and a small amount of coal for melting the lead in the
electrolytic method.
More important, however, is the greater saving of the metal values by
reason of increased yields of gold, silver, lead, antimony and bismuth,
and the freedom of the refined lead from bismuth.
Tables II, III, and IV show the composition of bullion, slimes and
refined lead.
Tables V, VI, VII, and VIII give the results obtained experimentally in
the laboratory on lots of a few pounds up to a few hundred pounds. The
results in Tables VI and VII were given me by the companies for which
the experiments were made.
TABLE I.—ANALYSES OF DROSS
For analyses of the lead from which this dross was taken, see Table II
───┬──────┬─────────┬─────────┬─────────┬─────────┬─────
│NO. IN│ │ │ │ │
NO.│TABLE │ CU. │ AS. │ SB. │ FE. │ZN.
│ II. │PER CENT.│PER CENT.│PER CENT.│PER CENT.│
───┼──────┼─────────┼─────────┼─────────┼─────────┼─────
1 │ 2 │ 0.0005 │ 0.0003 │ 0.0016 │ 0.0016 │none
2 │ 3 │ 0.0010 │ 0.0008 │ 0.0107 │ 0.0011 │ “
───┴──────┴─────────┴─────────┴─────────┴─────────┴─────
TABLE II.—ANALYSES OF BULLION
───┬─────────┬─────────┬─────────┬─────────┬────────
NO.│ FE. │ CU. │ SB. │ SN. │ AS.
│PER CENT.│PER CENT.│PER CENT.│PER CENT.│PER CENT.
───┼─────────┼─────────┼─────────┼─────────┼─────────
1 │ 0.0075 │ 0.1700 │ 0.5400 │ 0.0118 │ 0.1460
2 │ 0.0115 │ 0.1500 │ 0.6100 │ 0.0158 │ 0.0960
3 │ 0.0070 │ 0.1600 │ 0.4000 │ 0.0474 │ 0.1330
4 │ 0.0165 │ 0.1400 │ 0.7000 │ 0.0236 │ 0.3120
5 │ 0.0120 │ 0.1400 │ 0.8700 │ 0.0432 │ 0.2260
6 │ 0.0055 │ 0.1300 │ 0.7300 │ 0.0316 │ 0.1030
7 │ 0.0380 │ 0.3600 │ 0.4030 │ │ tr.
───┴─────────┴─────────┴─────────┴─────────┴─────────
───┬─────────┬─────────┬─────────┬─────────┬─────────
NO.│ AG. │ AU. │ PB. │ AG. │ AU.
│PER CENT.│PER CENT.│PER CENT.│OZ. P. T.│OZ. P. T.
───┼─────────┼─────────┼─────────┼─────────┼─────────
1 │ 1.0962 │ 0.0085 │ 98.0200 │ 319.7 │ 2.49
2 │ 1.2014 │ 0.0086 │ 97.9068 │ 350.4 │ 2.52
3 │ 1.0738 │ 0.0123 │ 98.1665 │ 313.2 │ 3.6
4 │ 0.8914 │ 0.0151 │ 97.9014 │ 260.0 │ 4.42
5 │ 0.6082 │ 0.0124 │ 98.0882 │ 177.4 │ 3.63
6 │ 0.6600 │ 0.0106 │ 98.2693 │ 192.5 │ 3.10
7 │ 0.7230 │ 0.0180 │ 98.4580 │ 210.9 │ 5.25
───┴─────────┴─────────┴─────────┴────────────────────
TABLE III.—ANALYSES OF SLIMES
─────────┬─────────┬─────────┬─────────┬─────────┬─────┬────┬─────
FE. │ CU. │ SB. │ SN. │ AS. │ PB. │ZN. │BI.
PER CENT.│PER CENT.│PER CENT.│PER CENT.│PER CENT.│ │ │
─────────┼─────────┼─────────┼─────────┼─────────┼─────┼────┼─────
1.27 │ 8.83 │ 27.10 │ 12.42 │ 28.15 │17.05│none│none
1.12 │ 22.36 │ 21.16 │ 5.40 │ 23.05 │10.62│ “ │ “
─────────┴─────────┴─────────┴─────────┴─────────┴─────┴────┴─────
TABLE IV.—ANALYSES OF REFINED LEAD
───┬───────┬───────┬───────┬───────┬──────┬───────┬──────┬──────┬─────
│ │ │ │ │ │ │ │ NI, │
│ CU. │ AS. │ SB. │ FE. │ ZN. │ SN. │ AG. │CO, CD│ BI.
NO.│ PER │ PER │ PER │ PER │ PER │ PER │ OZ. │ PER │ PER
│ CENT. │ CENT. │ CENT. │ CENT. │ CENT.│ CENT. │ P. T.│ CENT.│CENT.
───┼───────┼───────┼───────┼───────┼──────┼───────┼──────┼──────┼─────
1 │0.0006 │0.0008 │0.0005 │ │ │ │ │ │
2 │0.0003 │0.0002 │0.0010 │0.0010 │ none │ │ │ │
3 │0.0009 │0.0001 │0.0009 │0.0008 │ ” │ │ 0.24 │ │
4 │0.0016 │ │0.0017 │0.0014 │ │ │ 0.47 │ none │
5 │0.0003 │ │0.0060 │0.0003 │ │ │ 0.22 │ │
6 │0.0020 │ │0.0010 │0.0046 │ │ │ 0.22 │ none │
7 │0.0004 │ none │0.0066 │0.0013 │ none │0.0035 │ 0.14 │ │
8 │0.0004 │ │0.0038 │0.0004 │ ” │0.0035 │ 0.25 │ │
9 │0.0005 │ │0.0052 │0.0004 │ ” │0.0039 │ 0.28 │ │
10 │0.0003 │ none │0.0060 │0.0003 │ ” │0.0049 │ 0.43 │ │
11 │0.0003 │ ” │0.0042 │0.0013 │ ” │0.0059 │ 0.32 │ │
12 │0.0005 │ ” │0.0055 │0.0009 │ ” │0.0049 │ 0.22 │ │
13 │0.0005 │ ” │0.0055 │0.0007 │ ” │0.0091 │ 0.11 │ │
14 │0.0004 │ ” │0.0063 │0.0005 │ ” │0.0012 │ 0.14 │ │
15 │0.0003 │ ” │0.0072 │0.0003 │ ” │0.0024 │ 0.24 │ │
16 │0.0006 │ ” │0.0062 │0.0012 │ ” │0.0083 │ 0.22 │ │
17 │0.0006 │ ” │0.0072 │0.0011 │ │0.0080 │ 0.23 │ │
18 │0.0006 │ ” │0.0057 │0.0010 │ │0.0053 │ 0.34 │ │
19 │0.0005 │ ” │0.0066 │0.0016 │ │0.0140 │ 0.38 │ │
19 │0.0005 │ ” │0.0044 │0.0011 │ │0.0108 │ 0.35 │ │
20 │0.0004 │ ” │0.0047 │0.0015 │ │0.0072 │ 0.22 │ │
20 │0.0004 │ ” │0.0034 │0.0016 │ │ trace │ 0.23 │ │
21 │0.0022 │ ” │0.0010 │0.0046 │ none │0.0081 │ 0.38 │ none │ none
───┴───────┴───────┴───────┴───────┴──────┴───────┴───────────────────
TABLE V.—ANALYSES OF BULLION AND REFINED LEAD
──────────────┬───────────┬───────────┬───────────┬──────────
│ AG. │ CU. │ SB. │ PB.
│ PER CENT. │ PER CENT. │ PER CENT. │ PER CENT.
──────────────┼───────────┼───────────┼───────────┼───────────
Bullion │ 0.50 │ 0.31 │ 0.43 │ 98.76
Refined lead │ 0.0003 │ 0.0007 │ 0.0019 │ 99.9971
──────────────┴───────────┴───────────┴───────────┴───────────
TABLE VI.—ANALYSES OF BULLION AND REFINED LEAD
────────┬──────┬──────┬──────┬──────┬──────┬──────┬─────┬──────┬──────
│ CU. │ BI. │ AS. │ SB. │ AG. │ AG. │ AU. │ FE. │ ZN.
│ PER │ PER │ PER │ PER │ PER │ PER │ PER │ PER │ PER
│ CT. │ CT. │ CT. │ CT. │ CT. │ CT. │ CT. │ CT. │ CT.
────────┼──────┼──────┼──────┼──────┼──────┼──────┼─────┼──────┼──────
Bullion │0.75 │1.22 │0.936 │0.6832│358.89│ │1.71 │ │
Refined │ │ │ │ │ │ │ │ │
lead │0.0027│0.0037│0.0025│0.0000│ │0.0010│none │0.0022│0.0018
────────┴──────┴──────┴──────┴──────┴──────┴──────┴─────┴──────┴──────
TABLE VII.—ANALYSES OF BULLION, REFINED LEAD AND SLIMES
────────────┬─────┬──────┬───────┬───────┬───────┬───────┬──────┬─────
│ │ │ │ │ │ │FE,ZN,│
│ PB. │ CU. │ AS. │ SB. │ AG. │ AG. │NI,CO.│ BI.
│ PER │ PER │ PER │ PER │ OZ. │ PER │ PER │
│CENT.│ CENT.│ CENT. │ CENT. │ Per T.│ CENT. │ CENT.│
────────────┼─────┼──────┼───────┼───────┼───────┼───────┼──────┼─────
│ │ │ │ │about │ │ │
Bullion │96.73│0.096 │0.85 │ 1.42 │275[51]│ │ │
Refined lead│ │0.0013│0.00506│ 0.0028│ │0.00068│0.0027│trace
Slimes (dry │ │ │ │ │ │ │ │
sample) │ 9.05│1.9 │9.14 │29.51 │9366.9 │ │0.49 │trace
────────────┴─────┴──────┴───────┴───────┴───────┴───────┴──────┴─────
TABLE VIII.—ANALYSES OF BULLION, REFINED LEAD AND SLIMES
────────┬─────────┬─────────┬─────────┬─────────┬─────────┬────────
│ PB. │ CU. │ BI. │ AG. │ SB. │ AS.
│PER CENT.│PER CENT.│PER CENT.│PER CENT.│PER CENT.│PER CENT.
────────┼─────────┼─────────┼─────────┼─────────┼─────────┼─────────
Bullion │ 87.14 │ 1.40 │ 0.14 │ 0.64 │ 4.0 │ 7.4
Lead │ │ 0.0010 │ 0.0022 │ │ 0.0017 │ trace
Slimes │ 10.3 │ 9.3 │ 0.52 │ 4.7 │ 25.32 │ 44.58
────────┴─────────┴─────────┴─────────┴─────────┴─────────┴─────────
PART X
SMELTING WORKS AND REFINERIES
THE NEW SMELTER AT EL PASO, TEXAS
(April 19, 1902)
In July, 1901, the El Paso, Texas, plant of the Consolidated Kansas
City Smelting and Refining Company[52] was almost completely destroyed
by fire. The power plant, blast-furnace building and blast furnaces
were entirely destroyed, and portions of the other buildings were badly
damaged. The flames were hardly extinguished before steps were taken to
construct a new, modern and enlarged plant on the ruins of the old one,
and on April 15, 1902, nine months after the destruction of the former
plant, the new furnaces were blown in. In rebuilding it was decided to
locate the new power-house at some distance from the other buildings.
The furnaces have all been enlarged, each of the new lead furnaces (of
which there are seven) having about 200 tons daily capacity. These and
the three large copper furnaces have been located in a new position
in order to secure a larger building territory. The entire plant is
modern and up to date in every particular. One of the interesting
features is the substitution of crude oil as fuel in the boiler and
roasting departments. It is intended to use Beaumont petroleum for
the generation of power and the roasting of the ores instead of wood,
coal or coke, and it is expected that a considerable economy will be
effected by this means.
_Power Plant._—The power plant is complete in all respects. It is a
duplicate plant in every sense of the word, so that it will never be
necessary to shut the works down on account of the failure of any one
piece of machinery. There are seven boilers, having a total of 1250
h.p. The four blowers are unusually large, having a capacity of 30,000
cu. ft. of free air per minute. They are direct-connected to three
tandem compound condensing Corliss engines. No belts are used in this
plant, except for driving a small blower of 10,000 cu. ft. capacity,
which will act as a regulator. A large central electric plant has been
installed in the power-house, consisting of two direct-connected,
direct-current generators, mounted on the shafts of two cross-compound
condensing Nordberg-Corliss engines. The current from these generators
is transmitted through the plant, operating sampling works, briquetting
machinery, pumps, hoists, motors, cars, etc., displacing all the
small steam engines and steam pumps used in the old plant. The power
plant is provided with two systems for condensing; one being a large
Wheeler surface condenser, the other a Worthington central-elevated jet
condenser, the idea being to use the surface condenser during a short
period of the year when the water is so bad that it cannot be used in
the boilers. During the remainder of the year the jet condenser is in
service and the surface condenser can be cleaned. The condensed steam
from the surface condenser, with the necessary additional water, goes
back directly to the boilers when the surface condenser is in use. The
power-house is absolutely fireproof throughout, being of steel and
brick with iron and cement floors. It is provided with a traveling
crane, and no expense has been spared to make this, as all other
parts of the plant, complete in every respect. The main conductors
from the generators pass out through a tunnel into a brick and steel
lightning-arrester house, from which point the various distributing
lines go to different parts of the plant.
_Blast Furnaces._—There are seven large lead furnaces, each having a
capacity of 200 to 250 tons of charge per day, and three large copper
furnaces, each having a capacity of 250 to 300 tons per day. All of
the furnaces are enclosed in one steel fireproof building, the lead
furnaces being at one end and the copper furnaces at the other. Each
of the furnaces has its independent flue system and stack. An entirely
new system of feeding these furnaces has been devised, consisting of
a 6 ton charge car operated by means of a street railroad motor and
controller with third-rail system. The charge cars collect their charge
at the ore beds, lime-rock and coke storage, and are run on to 15 ton
hydraulic elevators. They are then elevated 38 ft. to the top of the
furnaces, traveling over them to the charging doors, through which the
loads are dumped directly into the furnaces. This system permits of two
men handling about 1000 tons per day. The same system and cars are used
for charging the copper furnaces, except that, as these furnaces are
much lower than the lead furnaces, the charge is dropped into a large
hopper, from which it is fed to the copper furnaces by a man on the
copper furnace feed-floor level.
NEW PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY AT MURRAY, UTAH
BY WALTER RENTON INGALLS
(June 28, 1902)
Murray is a few miles south of Salt Lake City, with which it is
connected by a trolley line. The new works are situated within a few
hundred yards of the terminus of the latter and in close juxtaposition
to the old Germania plant, which is the only one of the Salt Lake
lead-smelting works in operation at present. The new plant is of
special interest inasmuch as it is the latest construction for
silver-lead smelting in the United States, and may be considered as
embodying the best experience in that industry, the designers having
had access to the results attained at almost all of the previous
installations. It will be perceived, however, that there has been no
radical departure in the methods, and the novelties are rather in
details than in the general scheme.
The new works are built on level ground; there has been no attempt to
seek or utilize a sloping or a terraced surface, save immediately in
front of the blast furnaces, where there is a drop of several feet
from the furnace-house floor to the slag-yard level, affording room
for the large matte settling-boxes to stand under the slag spouts.
A lower terrace beyond the slag yard furnishes convenient dumping
ground. Otherwise the elevations required in the works are secured by
mechanical lifts, the ore, fluxes and coal being brought in almost
entirely by means of inclines and trestles.
The plant consists essentially of two parts, the roasting department
and the smelting department. The former comprises a crushing mill
and two furnace-houses, one equipped with Brückner furnaces and the
other with hand-raked reverberatories. The reverberatories are of
the standard design, but are noteworthy for the excellence of their
construction. Similar praise may be, indeed, extended to almost all
the other parts of the works, in which obviously no expense has been
spared on false grounds of economy. The roasting furnaces stand in a
long steel house; they are set at right angles to the longer axis of
the building, in the usual manner. At their feed end they communicate
with a large dust-settling flue, which leads to the main chimney of
the works. The ore is brought in on a tramway over the furnaces and is
charged into the furnaces through hoppers. The furnaces have roasting
hearths only. The fire-boxes are arranged with step-grates and closed
ash-pits, being fed through hoppers at the end of the furnace. The
coal is dumped close at hand from the railway cars, which are switched
in on a trestle parallel with the side of the building, which side is
not closed in. This, together with a large opening in the roof for
the whole length of the building, affords good light and ventilation.
The floor of the house is concrete. The roasted ore is dropped into
cars, which run on a sunken tramway passing under the furnaces. At the
end of this tramway there is an incline up which the cars are drawn
and afterward dumped into brick bins. From the latter it is spouted
into standard-gage railway cars, by which it is taken to the smelting
department. The roasted ore from the Brückner furnaces is handled in a
similar manner. The delivery of the coal and ore to the Brückners and
the general installation of the latter are analogous to the methods
employed in connection with the reverberatories.
The central feature of the smelting department is the blast-furnace
house, which comprises eight furnaces, each 48 by 160 in. at the
tuyeres. In their general design they are similar to those at the
Arkansas Valley works at Leadville. There are 10 tuyeres per side, a
tuyere passing through the middle of each jacket, the latter being
of cast iron and 16 in. in width; their length is 6 ft., which is
rather extraordinary. The furnaces are very high and are arranged for
mechanical charging, a rectangular brick down-take leading to the dust
chamber, which extends behind the furnace-house. The furnace-house is
erected entirely of steel, the upper floor being iron plates laid on
steel I-beams, while the upper terrace of the lower floor is also laid
with iron plates. As previously remarked, the lower floor drops down a
step in front of the furnaces, but there is an extension on each side
of every furnace, which affords the necessary access to the tap-hole.
The hight of the latter above the lower terrace leaves room for the
large matte settling-boxes, and the matte tapped from the latter runs
into pots on the ground level, dispensing with the inconvenient pits
that are to be seen at some of the older works. The construction of
the blast furnaces, which were built by the Denver Engineering Works
Company, is admirable in all respects. The eight furnaces stand in a
row, about 30 ft. apart, center to center. The main air and water pipes
are strung along behind the furnaces. The slag from the matte-settling
boxes overflows into single-bowl Nesmith pots, which are to be handled
by means of small locomotives. The foul slag is returned by means of a
continuous pan-conveyor to a brick-lined, cylindrical steel tank behind
the furnace-house, whence it is drawn off through chutes, as required
for recharging.
The charges are made up on the ground level, immediately behind the
furnace-house. The ore and flux are brought in on trestles, whence the
ore is unloaded into beds and the flux into elevated bins. These are
all in the open, there being only two small sheds where the charges are
made up and dumped into the cars which go to the furnaces. There are
two inclines to the latter. At the top of the inclines the cars are
landed on a transferring carriage by which they can be moved to any
furnace of the series.
The dust-flue extending behind the furnace-house is arranged to
discharge into cars on a tramway in the cut below the ground level.
This flue, which is of brick, connects with the main flues leading to
the chimney. The main flues are built of concrete, laid on a steel
frame in the usual manner, and are very large. For a certain distance
they are installed in triplicate; then they make a turn approximately
at right angles and two flues continue to the chimney. At the proper
points there are large dampers of steel plate, pivoted vertically, for
the purpose of cutting out such section of flue as it may be desired to
clean. Each flue has openings, ordinarily closed by steel doors, which
give access to the interior. The flues are simple tunnels, without
drift-walls or any other interruption than the arched passages which
extend transversely through them at certain places. The chimney is of
brick, circular in section, 20 ft. in diameter and 225 ft. high. This
is the only chimney of the works save those of the boiler-house.
The boiler-house is equipped with eight internally fired corrugated
fire-box boilers. They are arranged in two rows, face to face.
Between the rows there is an overhead coal bin, from which the coal is
drawn directly to the hoppers of the American stokers, with which the
boilers are provided. Adjoining the boiler-house is the engine-house;
the latter is a brick building, very commodious, light and airy. It
contains two cross-compound, horizontal Allis-Chalmers (Dickson)
blowing engines for the blast furnaces, and two direct-connected
electrical generating sets for the development of the power required
in various parts of the works. A traveling crane, built by the Whiting
Foundry Equipment Company, spans the engine-house. In close proximity
to the engine-house there is a well-equipped machine shop. Other
important buildings are the sampling mill and the flue-dust briquetting
mill.
A noteworthy feature of the new plant is the concrete paving, laid on a
bed of broken slag, which is used liberally about the ore-yard and in
other places where tramming is to be done. The roasting-furnace houses
are floored with the same material, which not only gives an admirably
smooth surface, but also is durable. The whole plant is well laid out
with service tramways and standard-gage spur tracks; the intention has
been, obviously, to save manual labor as much as possible.
THE MURRAY SMELTER, UTAH[53]
BY O. PUFAHL
(May 26, 1906)
This plant has been in operation since June, 1902. It gives employment
to 800 men. The monthly production consists of about 4000 tons of
work-lead and 700 tons of lead-copper matte (12 per cent. lead, 45 per
cent. copper). The work-lead is sent to the refinery at Omaha; the
matte to Pueblo, Colo. Most of the ores come from Utah; but in addition
some richer lead ores are obtained from Idaho, and some gold-bearing
ores from Nevada.
For sampling the Vezin apparatus is used, cutting out one-fifth in
each of three passes, crushing intervening, the sample from the third
machine being 1/625 of the original ore; after further comminution of
sample in a coffee-mill grinder, it is cut down further by hand, using
a riffle. The final sample is bucked down to pass an 80-mesh sieve, but
gold ores are put through a 120-mesh.
The steps in the smelting process are as follows: Roasting the poorer
ores in reverberatory furnaces and in Brückner cylinders. Smelting
raw and roasted ores, mixed, in water-jacketed blast furnaces,
for work-lead and lead-copper matte, the latter containing 15 per
cent. lead and 10 to 12 per cent. copper. Roasting the ground
matte, containing 22 per cent. of sulphur, down to ¾ per cent. in
reverberatory furnaces. Smelting the roasted matte together with acid
flux in the blast furnace for a matte with 45 per cent. copper and 12
per cent. lead.
Only the pyritic ores are roasted in Brückner furnaces, the lead ores
and matte being roasted in reverberatory furnaces. Each of the 20
Brückner furnaces, which constitute one battery, roasts 8 to 12 tons
of ore in 24 hours down to 5½ to 6 per cent. sulphur, with a coal
consumption of two tons. The charge weighs 24 tons. The furnaces make
one turn in 40 minutes. To increase the draft and the output, steam
at 40 lb. pressure is blown in through a pipe; this has, however,
resulted in increasing the quantity of flue dust to 10 to 15 per cent.
of the ore charged. Ten furnaces are attended by one workman with one
assistant, working in eight-hour shifts. For firing and withdrawing the
charge five men are required.
The gases from the Brückners and reverberatory furnaces pass into a
dust-flue 14 × 14 ft. in section and 600 ft. long, built of brickwork,
with concrete vault; in the stack (225 ft. high, 20 ft. diameter) they
unite with the shaft-furnace gases, the temperature of which is only 60
deg.
There are 12 reverberatory furnaces with hearths 60 ft. long and 16
ft. broad. They roast 14 tons of ore (or 13 tons of matte) in 24 hours
down to 3½ to 4 per cent. sulphur, consuming 32 to 34 per cent. of coal
figured on the weight of the charge. There are 12 working doors on each
side. The small coal (from Rock Springs, Wyoming), which is burnt on
flat grates, contains 5 per cent. ash and 3 to 5 per cent. moisture.
The roasted product is dumped through an opening in the hearth,
ordinarily kept closed with an iron plate, into cars which are raised
by electricity on a self-acting inclined plane. Their content is then
tipped over into a chute and cooled by sprinkling with water. From here
the roasted matte is conveyed to the blast furnace in 30-ton cars. The
roasted ore is tipped into the ore-bins.
There are eight blast furnaces, 48 × 160 in. at the tuyeres, of which
there are 10 on each of the long sides. The hight from the tuyeres to
the gas outlet is 20 ft., thence to the throat 6 ft.; the distance
of the tuyeres from the floor is 4 ft. The base is water-cooled. The
water-jackets of the furnace are 6 ft. high. The tuyeres (4 in.)
are provided with the Eilers automatic arrangement for preventing
the furnace gases entering the blast pipes. The blast pressure is
34 oz. The furnaces are furnished with the Arents lead wells; the
crucible holds about 30 tons of lead. The slag and the matte run into
a brick-lined forehearth (8 × 3 ft., 4 ft. deep), from which the slag
flows into pots holding 30 cu. ft., while the matte is tapped off into
flat round pans mounted on wheels.
The charge is conveyed to the feed-floor by electricity. The furnace
charge is 8000 lb. and 12 per cent. coke, with 30 per cent, (figured on
the weight of the charge) of “shells” (slag). Occasionally as much as
230 tons of the (moist) charge, exclusive of coke and slag, has been
handled by one furnace in 24 hours. During one month (September, 1904)
40,000 tons of charge were worked up, corresponding to a daily average
of 166 tons per furnace.
The lead in the charge runs from 13 to 14 per cent. on an average. The
limestone, which is added as flux, is quarried not far from the works.
The coke used is in part a very ordinary quality from Utah; in part a
better quality from the East, with 9 to 10 per cent. ash. The matte
amounts to 10 per cent. The slag contains 0.6 to 0.7 per cent. lead and
0.1 to 0.15 per cent. copper. The slag has approximately the following
composition: 36 per cent. silica, 23 per cent. iron (corresponding to
29.57 per cent. FeO), 23 per cent. lime, 3.8 per cent. zinc and 4 per
cent. alumina.
The work-lead is transferred while liquid from the furnaces to kettles
of 30 tons capacity, in which it is skimmed, and thence cast in molds
through a Steitz siphon. First, however, a 5.5 lb. sample is taken
out by means of a special ladle, and is cast into a plate. From this
samples of 0.5 a.t. are punched out at four points for the assay of the
precious metals.
For the purpose of precipitating the flue dust, the blast-furnace gases
are passed into brickwork chambers in which a plentiful deposition of
the heavier particles takes place. From here the gases go through an L
pipe of sheet iron, 18 ft. in diameter, to the Monier flues, which have
a cross-section of 256 sq. ft. and a total length of 2000 ft. A small
part of the flues is also built of brick. The gases unite with the hot
roaster gases just before entering the 225 ft. chimney. In the portion
of the blast-furnace dust first precipitated the silver runs 22 oz. per
ton, while that recovered nearer the stack contains only 8 oz. The flue
dust is briquetted with a small proportion of lime, and, after drying,
is returned to the blast furnaces.
THE PUEBLO LEAD SMELTERS[54]
BY O. PUFAHL
(May 12, 1906)
At the Pueblo plant, ores containing over 10 per cent. lead are not
roasted, but are added raw to the charge. For such material as requires
roasting there are in use five Brückner furnaces. The charge is 24 tons
for 48 to 60 hours; the furnaces make one revolution per minute and
roast the ore down to 6 per cent. sulphur. There are also two O’Harra
furnaces, each roasting 25 tons daily, and 10 reverberatory furnaces 75
ft. in length, each roasting 15 tons of ore daily down to 4 per cent.
sulphur.
The charge for smelting is prepared from roasted ore, together with
Idaho lead ore, Cripple Creek gold ore, briquetted flue dust, slag
and limestone. There are seven water-jacketed furnaces, which smelt,
each, 150 tons of charge per day. The furnaces have 18 tuyeres, blast
pressure 34 oz., cross-section at the tuyeres 48 × 148 in. They are
charged mechanically by a car of 4 tons’ capacity.
The output of lead is 11 to 15 tons per furnace. The matte, which
is produced in small quantity, contains 8 to 12 per cent. lead and
the same percentage of copper. It is crushed by rolls, roasted in
reverberatory furnaces, and smelted with ores rich in silica. The matte
resulting at this stage, running 45 to 50 per cent. in copper, is
shipped to be further worked up for blister copper.
The work-lead is purified by remelting in iron kettles, the cupriferous
dross being pressed dry in a Howard press, and sent to the blast
furnaces. The work-lead is sent to the refineries at Omaha, Neb., or
Perth Amboy, N. J.
To collect the flue dust the waste gases are passed through long brick
flues. The chimneys are 150 to 200 ft. high, and 15 ft. in diameter.
They stand 75 ft. above the ground level of the blast furnaces. The
comparatively small proportion of flue dust produced (0.9 per cent. of
the charge) is briquetted, together with fine ore and 5 per cent. of
a thick paste of lime. For this purpose a White press is used, which
makes six briquets at a time, and handles 10 tons per hour.
According to a tabulation of the results of five months’ running, the
proportion of flue dust at several works of the American Smelting and
Refining Company is as follows:
Globe Plant, Denver 0.5% of the charge.
Pueblo Plant, Pueblo 0.9% ” ” ”
Eilers’ Plant, Pueblo 0.5% ” ” ”
East Helena Plant, Helena 0.3% ” ” ”
Arkansas Valley Plant, Leadville 0.2% ” ” ”
Murray Plant, Murray, Utah 1.2% ” ” ”
The fuel used is of very moderate quality. The coke (from beehive
ovens) carries up to 17 per cent. ash, the coal 10 to 18 per cent. The
monthly production is 2300 tons of work-lead and 150 tons of copper
matte (45 to 50 per cent. copper).
At the Eilers plant all sulphide ores, except the rich Idaho ore, are
roasted down to 5 to 7 per cent. S in 15 reverberatory furnaces, 60 to
70 ft. in length, each furnace roasting 15 tons per 24 hours, in six
charges.
The flue dust is briquetted together with fine Cripple Creek ore,
pyrites cinder from Argentine, Kan., Creede ores rich in silica and
10 per cent. lime. The residue from the zinc smeltery (U. S. Zinc
Company), which is brought to this plant (600 tons a month containing
nearly 10 per cent. lead), is taken direct to the blast furnaces.
Of the latter there are six, each with 18 tuyeres, which handle per
24 hours 160 to 180 tons of charge, containing on an average 10 per
cent. of lead in the ore, with 10 per cent. of coke, figured on the
charge. The average monthly production of a furnace is about 360 tons
of work-lead, which is purified at the Pueblo plant. The furnaces
are charged by hand. Of the slag, 30 per cent., as shells, etc., is
returned to the charge. The monthly production of work-lead is 2000
tons, carrying 150 oz. of silver and 2 to 6 oz. of gold per ton.
The matte amounts to about 8.3 per cent., and contains 12 per cent.
copper. It is concentrated up to 45 per cent. Cu, which is shipped (150
tons a month) for smelting to blister copper.
THE PERTH AMBOY PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY[55]
BY O. PUFAHL
(January 27, 1906)
These works were erected in 1895 by the Guggenheim Smelting Company.
They are situated on Raritan Bay, opposite the southern point of Staten
Island, in a position offering excellent facilities for transportation
by land and by water. The materials worked up are base lead bullion
and crude copper, containing silver and gold, chiefly drawn from the
company’s smelteries in the United States and Mexico. Silver ore is
received from South America. The ores and base metals from Mexico and
South America are brought to Perth Amboy by the company’s steamships
(American Smelters Steamship Company).
_Ore Smelting._—The silver ore from South America (containing antimony
and much silver, together with galena, iron and copper pyrites) is
crushed by rolls and is roasted down from 26 per cent. to 3 per cent.
S in 11 reverberatory furnaces, 70 ft. long and 15 ft. wide (inside
dimensions). It is then mixed with rich galena from Idaho, pyrites
cinder, litharge, copper skimmings, and residues from the desilverizing
process, together with limestone, and is smelted for work-lead and
lead-copper matte in three water-jacketed furnaces, using 12 per cent.
coke, figured on the ore in the charge. Of these furnaces one has 12
tuyeres; it measures 42 × 96 in. in cross-section at the tuyeres, and
6 ft. 3 in. by 8 ft. at the charging level. The hight of charge is 16
ft. The other two furnaces have 16 tuyeres each, their cross-section at
the tuyeres being 44 in. by 128 in., at the charging level 6 ft. 6 in.
by 12 ft., and hight of charge 16 ft. The furnaces are operated at a
blast pressure of 35 oz. per square inch. The temperature of the gases
at the throat is 140 deg. F. (60 deg. C.) measured with a Columbia
recording thermometer, which works very well. These furnaces reduce,
respectively, 100 to 120 and 130 to 140 tons of charge per 24 hours;
they are also used for concentrating roasted matte.
_Copper Refining._—The crude copper is melted in two furnaces of 125
tons aggregate daily capacity, and is molded into anodes by Walker
casting machines. Twenty-six anodes are lifted out of the cooling
vessel at a time, and are taken to the electrolytic plant.
The electrolytic plant comprises two systems, each of 408 vats. The
current is furnished by two dynamos, each giving 4700 amperes at 105
volts. The cathodes remain in the bath for 14 days. The weight of the
residual anodes is 15 per cent.
The anode mud is swilled down into reservoirs in the cellar as at
Chrome (De Lamar Copper Refining Company), is cleaned, dried and
refined in a similar manner.
For melting the cathodes there are two reverberatory furnaces of
capacity for 75 tons per 24 hours. The wire-bars and ingots are cast
with a Walker machine. About 3200 tons of refined copper are produced
per month.
_Copper Sulphate Manufacture._—The lyes withdrawn from the electrolytic
process are worked up into copper sulphate, shot copper being added.
This latter is prepared in a reverberatory furnace from matte obtained
as a by-product in working up the lead. About 200 tons of copper
sulphate are thus produced per month; the process used is the same as
at the Oker works. Lower Harz, Germany. The crystals are rinsed, dried
and packed in strong wooden barrels.
_Lead Refining._—The working up of the Mexican raw lead is carried
out under the supervision of the customs officers. The lead, which is
imported duty free, must be exported again. From each bar a sample is
cut from above and below by means of a punch entering half way into the
bar. For refining the lead there are four reverberatory furnaces of 60
tons capacity, with hearths 17 ft. 9 in. by 12 ft. 6 in., a mean depth
of 14 in., and a grate area of 2 ft. 6 in. by 6 ft.; in addition to
these there is a furnace of 80 tons capacity with a hearth 19 ft. 7½
in. by 9 ft. 6 in., a mean depth of 18 in., and grate area of 3 ft. by
6 ft.
For desilverizing the softened lead there are five kettles, each of
60 tons capacity, 10 ft. 3 in. diameter and 39 in. depth. The zinc
is stirred in with a Howard mechanical stirrer and the zinc scum is
pressed dry in a Howard press, which gives a very dry scum. The latter
is then, while still warm, readily hammered into pieces for the retorts.
The desilverized lead is refined in five reverberatory furnaces, of
which four take a charge of 50 tons each, and one of 65 tons. The
production of desilverized lead is 5000 to 5500 tons a month.
The distillation of the zinc crusts is carried out in 18 oil-fired
Faber du Faur tilting furnaces. Each retort receives a charge of
1200 lb. of broken-up crust and a little charcoal. The distillation
lasts 6 to 7 hours. Fifty gallons of petroleum residues are consumed
per charge. The oil is blown into the furnace with a compressed
air atomizer. After withdrawing the condenser, which runs on a
traveling support, the argentiferous lead is poured directly from
the tilted retort into an English cupel furnace. Seven such furnaces
(magnesia-lined, with movable test) are in use, of which each works
up 4.5 to 5 tons of retort metal in 24 hours. The furnaces are
water-jacketed. The blast is introduced by the aid of a jet of steam.
Three tons of coal are used per 24 hours.
_Gold and Silver Parting._—The doré bars are cast into anodes for
electrolytic parting by the Moebius process. The plant consists of 144
cells in 24 divisions. The mean composition of the electrolytic bath is
said to be as follows: 10 per cent. free nitric acid, 17 grams silver,
and 35 to 40 grams copper per liter. The current is furnished by a 62
k.w. dynamo. One cell consumes 260 amp. at 1.75 volts. One k.w. gives
a yield of 1600 oz. fine silver per 24 hours. The daily production
of silver is almost 100,000 oz., and is exceeded at no other works.
About $3,000,000 worth of metal is always on hand in the different
departments.
THE NATIONAL PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY[56]
BY O. PUFAHL
(April 14, 1906)
This plant, at South Chicago, Ill., refines base lead bullion. It
comprises four reverberatory furnaces, of which one takes a charge of
100 tons, one 80 tons, and the other two 30 tons each; one of the small
furnaces is being torn down, and a 120 ton furnace is to be built in
its place. The furnaces are fired with coal from Southern Illinois,
which contains 11 per cent. of ash.
In softening the bullion, the time for each charge is 10 hours. The
first portion tapped consists of dross rich in copper, which is
followed by antimonial skimmings and litharge.
The copper dross is melted up in a small reverberatory furnace,
together with galena from Wisconsin (containing 80 per cent. lead),
for work-lead and lead-copper matte, the latter containing about 35
per cent. of copper; this matte is enriched to 55 per cent. copper
by the addition of roasted matte, and is finally worked up for crude
copper (95 per cent.) in a reverberatory furnace. All the copper so
produced is used in the parting process for precipitating the silver.
The antimonial skimmings are smelted in a reverberatory furnace,
together with coke cinder, for lead and a slag rich in antimony, which
is reduced to hard lead (27 per cent. antimony, 0.5 per cent. copper,
0.5 per cent. arsenic) in a small blast furnace, 14 ft. high, which has
8 tuyeres.
The softened lead is tapped off into cast-iron desilverizing pots,
which usually outlive 200 charges; in isolated cases as many as
300. For desilverizing, zinc from Pueblo, Colo., is added in two
instalments, being mixed in by means of a Howard stirrer. After the
first addition there remains in the lead 7 oz. of silver per ton;
after the second only 0.2 oz. The first scum is pressed in a Howard
press and distilled; the second is ladled off and is added to the next
charge. The Howard stirrer is driven by a small steam engine suspended
over the kettle; the Howard press by compressed air.
For distilling zinc scum, 12 Faber du Faur tilting retorts, heated with
petroleum residue, are used. The argentiferous lead (with 9.6 per cent.
silver) is transferred from the retort to a pan lined with refractory
brick, which is wheeled to the cupelling hearth and raised by means of
compressed-air cylinders, so as to empty its molten contents through a
short gutter upon the cupelling hearth. The cupelling hearths are of
the water-cooled English type, and are heated by coal with under-grate
blast. The cast-iron test rings, with reinforcing ribs, are made in two
pieces, slightly arched and water-cooled; they are rectangular, with
rounded corners, and are mounted on wheels. The material of the hearth
is marl.
Argentiferous lead is added as the operation proceeds, and finally the
doré bullion is poured from the tilted test into thick bars (1100 oz.)
for parting.
The desilverized lead is refined in charges of 28 tons (4 to 5 hours)
and 80 to 90 tons (8 to 10 hours), introducing steam through four to
eight half-inch iron pipes. The first skimmings contain a considerable
proportion of antimony and are therefore added to the charge when
reducing the antimonial slags in the blast furnace. The litharge is
worked up in a reverberatory furnace for lead of second quality. The
refined lead is tapped off into a kettle, from which it is cast into
bars through a siphon.
The parting of the doré bullion is carried out in tanks of gray cast
iron, in which the solution is effected with sulphuric acid of 60 deg.
B. The acid of 40 deg. B. condensed from the vapors is brought up to
strength in leaden pans. In a second larger tank, which is slightly
warmed, a little gold deposits from the acid solution of sulphates.
The solution is then transferred (by the aid of compressed air) to the
large precipitating tank, and diluted with water. It is here heated
with steam, and the silver is rapidly precipitated by copper plates
(125 plates 18 × 8 × 1 in.) suspended in the solution from iron hooks
covered with hard lead. After the precipitation, the vitriol lye is
siphoned off, the silver is washed in a vat provided with a false
bottom, is removed with a wooden shovel, and is pressed into cakes 10 ×
10 × 6 in.
The refining is finished on a cupelling hearth fired with petroleum
residue, adding saltpeter, and removing the slag by means of powdered
brick. After drawing the last portion of slag the silver (0.999 fine)
is kept fused under a layer of wood-charcoal for 20 minutes, and is
then cast into iron molds, previously blackened with a petroleum flame.
The bars weigh about 1100 oz.
The gold is boiled with several fresh portions of acid, is washed and
dried, and finally melted up with a little soda in a graphite crucible.
It is 0.995 fine.
The lye from the silver precipitation, after clearing, is evaporated
down to 40 deg. B. in leaden pans by means of steam coils, and is
transferred to crystallizing vats. The first product is dissolved
in water, the solution is brought up to 40 deg. B. strength, and is
allowed to crystallize. The purer crystals so obtained are crushed, and
are washed and dried in centrifugal apparatus; they are then sifted and
packed in wooden casks in two grades according to the size of grain.
The very fine material goes back into the vats. From the first strongly
acid mother liquor, acid of 60 deg. B. is prepared by concentrating in
leaden pans, and this is used for the parting operation.
THE EAST HELENA PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY[57]
BY O. PUFAHL
(April 28, 1906)
The monthly production of these works is about 1500 tons of base
bullion (containing 150 oz. Ag and 4 to 6 oz. Au per ton), and 200 tons
of 45 per cent. copper matte. The base bullion is shipped to South
Chicago, the matte to Pueblo.
The ore-roasting is done in two batteries of eight reverberatory
furnaces and 16 Brückner furnaces, the resulting product containing on
an average 20 per cent. lead and 3 per cent. sulphur. The charge for
the blast furnaces consists of roasted ore, rich galena, argentiferous
red hematite, briquetted flue dust, slag (shells) from the furnace
itself, lead skimmings, scrap iron and limestone.
Four tons of the charge are dumped over a roller into a low car, which
is then drawn up an inclined plane to the charging gallery by an
electric motor and is then dumped into the furnace.
The two rectangular blast furnaces (Eilers’ type) have eight tuyeres on
each of their longer sides and cast-iron water-jackets of 6 ft. hight.
The blast is delivered at a pressure of 40 oz. The lead is drawn off
through a siphon tap into a cooling kettle. The furnace has a large
forehearth for separating the matte and the slag. The slag is received
by a two-pot Nesmith truck, having an aggregate capacity of 14 cu. ft.
These trucks are hauled to the dump by an electric locomotive. The
shells are returned to the furnace with the charge.
The matte (with about 6 per cent. Cu and the same percentage of lead)
is tapped off into iron molds and after cooling is crushed to 0.25-in.
size, to be roasted in the reverberatory furnaces and smelted up
together with roasted ore for a 15 per cent. matte. The latter is
crushed, roasted and separately smelted together with silicious ore
for 45 per cent. matte, which is then sent to Pueblo to be worked up
into blister copper. The small quantity of speiss which is formed is
broken up and returned to the blast furnaces with the charge. The slag
contains 0.5 to 0.8 per cent. lead and 0.5 oz. silver per ton.
THE GLOBE PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY[58]
BY O. PUFAHL
(May 5, 1905)
This plant produces 1800 tons of base bullion per month and 200 tons
of lead-copper matte containing 45 to 52 per cent. of copper. The ores
smelted are mostly from Colorado, but include also galena from the Cœur
d’Alene and other supplies. The limestone is quarried 14 miles from
Denver; coke and coal are brought from Trinidad, Colo.
All sulphides, except the slimes, concentrates and the rich Idaho ores,
are roasted. For roasting there are:
(1) Fifteen reverberatory furnaces, five of which measure 60 × 14 ft.,
and the other ten 80 × 16 ft. externally. In 24 hours these roast six
charges of 4400 lb. (average) of moist ore (2.15 tons of dry ore) from
28 to 30 per cent. down to 3 to 4 per cent. sulphur. Each furnace is
attended by three men working in 12-hour shifts; the stoker earns
$2.75; the roasters, $2.30.
(2) Two Brown-O’Harra furnaces, 90 ft. long, with two hearths, and a
small sintering furnace attached. They have three grates on each long
side, and each roasts 26 tons of ore in 24 hours down to ¾ per cent.
sulphur.
(3) Twelve Brückner furnaces, each taking 24 tons’ charge, with
under-grate blast, the air being fed into the cylinders by a steam jet.
According to the zinc content of the ores the roasting operation lasts
70 to 90 hours, the furnace making one revolution per hour. The roasted
product from the Brückner furnaces is pressed into briquets, together
with fine ore, flue dust and lime.
The smelting is carried out in seven blast furnaces, with 16 tuyeres,
blast at 2-lb. pressure, hight of furnace 18 ft. 6 in., section at
the tuyeres 42 × 144 in. The charge is 120 to 150 tons exclusive of
slag and coke. The slag and the matte are tapped off together into
double-bowl Nesmith cars, which are hauled, by an electric locomotive,
to a reverberatory furnace (hearth 20 × 12 ft.) in which they are kept
liquid, for several hours, in charges of 14 to 15 tons, in order to
effect complete separation. A little work-lead is obtained in this
operation, while the matte is tapped off into cast-iron pans of one
ton capacity, and the slag, 0.5 to 0.6 per cent. lead, 0.6 to 0.7 oz.
silver, is removed in 5-ton pots to the dump.
The matte is broken up, crushed to 0.25 in. size, roasted in the
reverberatory furnaces, smelted for a 45 to 52 per cent. copper matte,
which is shipped to be further worked up into blister copper. The crude
matte contains 10 to 12 per cent. copper, 12 to 15 per cent. lead, 40
oz. silver and 0.05 oz. gold.
From the siphon taps of the blast furnaces the work-lead is transferred
to a cast-iron kettle of 33 tons’ capacity. Here the copper dross
is removed, the metal is mixed by introducing steam for 10 minutes,
sampled, and the lead is cast into bars through siphons. It contains
about 2 per cent. antimony, 200 oz. silver and 8 oz. gold. This product
is refined at Omaha.
The blast-furnace gases pass through a flue 1200 ft. long, and enter
the bag-house, in which they are filtered through 4000 cotton bags 30
ft. long and 18 in. in diameter. These bags are shaken every 6 hours.
The material which falls to the floor is burnt where it lies, sintered
and returned to the blast furnaces.
In the engine house there are four Connersville blowers, two of which
are No. 8 and two of No. 7 size. Each blast furnace requires 45,000 cu.
ft. of air a minute.
The works give employment to 450 men, whose wages (for 10-to 12-hour
shifts) are $2 to $3.
LEAD SMELTING IN SPAIN
BY HJALMAR ERIKSSON
(November 14, 1903)
A few notes, gathered during a couple of years while I was employed
at one of the large lead works in the southeastern part of Spain, are
of interest, not as showing good work, but for comparing the results
obtained in modern practice with those obtained by what is probably the
most primitive kind of smelting to be found today. The plant about to
be described may serve as a general type for that country. As far as I
know, the exceptions are a large plant at Mazarron, fully up to date
and equipped with the most modern improvements in every line; a smaller
plant at Almeria, also in good shape, and the reverberatory smelting of
the carbonates at Linares. It should be kept in mind, however, that the
conditions are peculiar, iron and machinery being very expensive and
manual labor very cheap.
[Illustration: FIG. 41.—Spanish Lead Blast Furnace.]
About 4 ft. above the tuyeres the furnace is built of uncalcined brick
made of a black graphitic clay found in the mines near by; the upper
part is of common red brick. The entire cost of one furnace does not
reach $100. The flue leads to a main gallery 3.5 by 7 ft., which goes
down to the ground, and extends several times around a hill, the
chimney being placed on the top of the hill, considerably above the
furnace level. The gallery is about 10,000 ft. long, and is laid down
in the earth, with the arched roof just emerging. It is all built of
rough stone, the inside being plastered with gypsum. The furnace has
three tuyeres of 3 in. diameter. The blast pressure is generally 4 to
6 in. of water. Neither feeding floor nor elevators are used, only a
couple of scaffolds, the charge being lifted up gradually by hand in
small convenient buckets made of sea-grass. When charging the furnace,
coke is piled up in the center, and the mixture of ore, fluxes and slag
is charged around the walls. The slag and matte are left to run out
together on an inclined sand-bed. The matte, flowing more quickly, goes
further and leaves the slag behind, but the separation thus obtained
is, of course, very unsatisfactory. The charge mixture is weighed and
made for each furnace every morning. When it is all put through, the
furnace is run down very low, without any protecting cover on the top;
several iron bars are driven through the furnace at the slag-tap level,
for holding up the charge; the lead is all tapped out; a big hole is
made in the crucible for the purpose of cleaning it out; all accretions
are loosened with a bar; the hole is closed with mud of the graphitic
clay; bars are removed, when the crucible is filled with coke from the
center and the charging is continued. In this way a furnace can be kept
running for any length of time, but at a great loss of heat, and with a
great increase of flue dust.
The current practice, in many parts of Spain, is to run the same number
of ore-smelting and of matte-smelting furnaces. All the slag and the
raw matte, produced by the ore-smelting furnaces, is re-smelted in the
matte furnaces, together with some dry silver ores. No lead at all is
produced in the matte furnaces, only a matte containing up to 150 oz.
silver per ton and 25 to 35 per cent. of the lead charged on them. This
rich matte is calcined in kilns, and smelted together with the ore
charge.
The ores we smelted were galena ranging from 5 to 83 per cent. lead
and about 250 oz. silver per ton of lead; dry silver ores containing
up to 120 oz. silver per ton, and enough of the Linares carbonates for
keeping the silver below 120 oz. per ton in the lead. The gangue of the
galena was mainly iron carbonate. Most of that ore was hand picked and
of nut size. Machine concentrates with more than 30 per cent. lead or
containing much pyrite were calcined; everything else was smelted raw.
The flux exclusively used, before I came, was carbonate of iron, which,
by the way, was considered a “cure-for-all.” The slag analyses showed:
CaO, below 4 per cent.
FeO, above 45 per cent.
SiO₂, about 30 per cent.
BaO, 5 to 10 per cent.
Al₂O₃, 5 to 10 per cent.
Pb, by fire assay, 0.75 to 2.5 per cent.
Ag, by fire assay, 2 to 3 oz. per ton.
The specific gravity of the slag was about 5, or practically the same
as that of the matte. The output of metallic lead was about 70 per
cent.; of silver, 84 per cent. The working hight of the furnaces—tuyere
level to top of charge—was at that time only 7 ft., and I was told that
it had been still lower before.
To the working hight of the furnaces was added 2 ft., simply by putting
up the charging doors that much. A very good limestone was found just
outside the fence around the plant. Enough limestone was substituted
for the iron carbonate, to keep the lime up to 12 per cent. in the
slag, reducing the FeO to below 35 per cent. and the specific gravity
to below four.
The result of these alterations was an increase in the output of
metallic lead, from 76 to 85 per cent.; of silver from 84 to 90 per
cent.; a comparatively good separation of slag and matte, and a slag
running about 0.5 to 0.75 per cent. Pb and 1.5 oz. Ag per ton.
Owing to the great extent of the gallery, and the consequent good
condensation of the flue dust, the total loss of lead and silver was
much smaller than would be expected; in no case being found above 4 per
cent.
The composition of the charge was 55 per cent. ore and roasted matte,
13 per cent. fluxes, and 32 per cent. slag. Coke used was 11 per cent.
on charge, or 20 per cent. on ore smelted. Each furnace put through 10
to 15 tons of charge, or 7 tons of ore, in 24 hours. Eight men and two
boys were required for each furnace, including slag handling and making
up of the charge. The cost of smelting was 17 pesetas per ton of ore,
which at the usual premium (£1 = 34 pesetas = $4.85) equals $2.43. This
cost is divided as follows:
Coke $1.47
Fluxes 0.04
Labor 0.65
Coal for power 0.10
General expenses 0.17
————-
Total $2.43
This $2.43 per ton includes all expenses of whatever kind. The iron
carbonate flux contained lead and silver, which was not paid for. The
fluxes are credited for the actual value of this lead and silver.
Without making this discount, the cost of flux would amount to 26c. per
ton, making the entire smelting cost come to $2.65. As an explanation
of the low cost of labor, it may be noted that the wages were, for
the furnace-man, 2.25 pesetas, or 32c. a day; for the helpers, 1.75
pesetas, or 25c. a day.
The basis for purchasing the galena ore may here be given, reduced to
American money; lead and silver are paid for according to the latest
quotations for refined metals given by the _Revista Minera_, published
at Cartagena. (The quotations are the actual value in Cartagena of the
London quotations.)
The following discounts are made: 5 per cent. for both silver and
lead; $6.40 per ton on ore containing 7 per cent. Pb and below; this
rises gradually to a discount of $7.75 per ton of ore containing 30 per
cent. Pb and above.
The transportation is paid by the purchaser and amounts to about $1.20
per ton of ore.
The dry silver ores were cheaper than this and the lead carbonates much
more expensive.
LEAD SMELTING AT MONTEPONI, SARDINIA[59]
BY ERMINIO FERRARIS
(October 28, 1905)
In dressing mixed lead and zinc carbonate ores by the old method of
gradual crushing with rolls, middling products were obtained, which
could be further separated only with much loss. Inasmuch as the losses
in the metallurgical treatment of such mixed ore were reckoned to be
less than in ore dressing, these between-products at Monteponi were
saved for a number of years, until there should be enough raw material
to warrant the erection of a small lead and zinc smeltery.
In 1894 the lead smeltery in Monteponi was put in operation; in 1899
the zinc smeltery was started. At about the same time the reserves of
lead ore were exhausted, and the lead plant then began to treat all the
Monteponi ores and a part of those from neighboring mines.
As will be seen from the plan (Fig. 42), the smelting works cluster
in terraces around the mine shaft, covering an area of about 3000 sq.
m. (0.75 acre); the ore stocks and the pottery of the zinc works are
located in separate buildings.
During the first years of working, the slag had purposely been kept
very rich in zinc, in the hope of utilizing it later for the production
of zinc oxide. It had an average zinc content of 16.80 per cent., or 21
per cent. of zinc oxide, with about 32 per cent. SiO₂, 25 per cent.
FeO, and 14 per cent. lime. According to the recent experiments, this
slag can very well be used for oxide manufacture, in connection with
calamine rich in iron. The slag made at the present time has only 15
per cent. ZnO; 25 per cent. SiO₂; 16 per cent. CaO; 3 per cent. MgO;
33 per cent. FeO; 2.5 per cent. Al₂O₃, and 2 per cent. BaO, and
small quantities of alkalies, sulphur and lead (1 to 1.5 per cent).
The following classes of ore are produced at Monteponi:
1. Lead carbonates, with a little zinc oxide; these ores are screened
down to 10 mm. The portion held back by the screen is sent straight
to the shaft furnaces; the portion passing through is either roasted
together with lead sulphides, or is sintered by itself, according to
circumstances.
2. Dry lead ores, mostly quartz, with 10 to 15 per cent. lead, which
are mixed for smelting with the lead carbonates.
[Illustration: FIG. 42.—General Plan of Works.]
3. Lead sulphides, which are crushed fine and roasted dead. Quartz
sand is added in the roasting, in order to decompose the lead sulphate
and produce a readily fusible silicate; as quartz flux, fine sand from
the dunes on the coast is used. This is a product of decomposition of
trachyte, and contains 88 per cent. of silica, together with alkalies
and alumina. The roast is effected in two hand-raked reverberatory
furnaces, 18 m. long, which turn out 12,000 kg. of roasted ore in
24 hours, consuming 1800 kg. of English cannel coal, or 2400 kg.
of Sardinian lignite. There is also a third reverberatory furnace,
provided with a fusion chamber, which is used for roasting matte and
for liquating various secondary products.
The charge for the shaft furnace, as a rule, consists of 50 per cent.
ore (crude and roasted), 20 per cent. fluxes and 30 per cent. slag
of suitable origin. The fluxes used are limestone from the mine,
containing 98 per cent. CaCO₃, and limonite from the calamine
deposits. This iron ore contains 48 per cent. Fe, not more than 4 per
cent. Zn, a little lead and traces of copper and silver.
A shaft furnace will work up a charge of 60 tons, equal to 30 tons of
ore, in 24 hours, with a coke consumption of 12 per cent. of the weight
of the charge and a blast pressure of 50 mm. of mercury. There are
three furnaces, of which two are used alternately for smelting lead
ores, while one smaller furnace serves for smelting down products, such
as hard lead, copper matte and copper bottoms.
[Illustration: FIG. 43.—Elevation of works on line A B C D E F of Fig.
42.]
Figs. 43 to 46 show one of the furnaces. It will be seen at once that
its construction is similar to that of the standard American furnaces.
Pilz furnaces were tried in the first few years, but were finally
abandoned, as they could not be kept running for any satisfactory
length of time with slags rich in zinc. Diluting the slag, on the other
hand, would have led to an increased coke consumption, and would have
rendered the slag itself worthless. The furnace, however, differs in
several respects from its American prototype; the following are some of
the chief characteristics peculiar to it:
[Illustration: Section E F. Section G H.
FIG. 44.—Shaft Furnace for Lead Smelting.]
The chimney above the feed-floor covers one-third of the furnace
shaft, and is turned down in the form of a siphon, to connect with
the flue-dust chamber. The lateral faces, which are left open, serve
as charging apertures; the central one of these, provided with a
counterbalanced sheet-iron door, is used for charging from cars. The
square openings at the ends, which are covered with cast-iron plates,
are used for barring down the furnace shaft and may also be used for
charging. By this arrangement, together with the two hoppers placed
laterally on the chimney, it is possible to distribute the charge in
any desired manner over the whole cross-section of the furnace. This
arrangement greatly facilitates the removal of any accretions in the
furnace shaft, as the centrally placed chimney catches all the smoke,
while the charge-holes render the furnace accessible on all sides.
In case of large accretions being formed, the whole furnace can be
emptied, cleaned and restarted in 24 to 36 hours.
The smelting cone is enclosed by cast-steel plates 50 cm. high, instead
of having a water-jacket. These are cooled as desired by turning a
jet of water on them. The plates are connected to the furnace shaft
by a bosh wall 25 cm. thick, which is surrounded with a boiler-plate
jacket. These jacket plates also are cooled from the outside by sprays
of water. With this arrangement the consumption of water is less than
with water-jackets, as a part of the water is vaporized, and the danger
of leakage of the jackets is avoided. The cast-steel plates are made
in two patterns; there are two similar side-plates, each with four
slits for the tuyeres, and two end-plates, provided with a circular
breast of 30 cm. aperture, for tapping the slag. The breast is cooled
by water flowing down, and is closed in front by a plate of sheet
iron, in which is the tap-hole for running off the slag. When cleaning
out, this sheet-iron plate is removed and the breast is opened, thus
providing easy access to the hearth. The four cast-steel plates are
anchored together with bolts at their outer ribs, and rest on two long,
gutter-shaped pieces of sheet iron, which carry off all the water which
flows down, and keep it away from the brickwork of the hearth.
[Illustration: Section J L. Section C D.
FIG. 45.—Shaft Furnace.]
The hearth, cased with boiler plate and rails, has at the side a
cast-iron pipe of 10 cm. diameter for drawing off the lead to the
outside kettle; this pipe has a slight downward inclination, to prevent
the slag flowing out; every 20 minutes lead is tapped, and the end of
the pipe is then plugged up with clay.
The furnace shaft is supported upon a hollow mantel, which serves at
the same time as blast-pipe. The blast-pipe has eight lateral tees,
which are connected by canvas hose with the eight tuyeres. The mouth
of the tuyeres has the form of a horizontal slit, whereby the air is
distributed more evenly over the entire zone of fusion.
[Illustration: FIG. 46.—Shaft Furnace for Lead Smelting. (Section A B.)]
The precipitation of flue dust is effected in a brick condensing
chamber, placed near the beginning of the main flue. The main flue
terminates on the hill (see Fig. 43) in a chimney, the top of which
is 160 m. above the ground level of the works, affording excellent
draft. The condensing chamber (Figs. 49 to 51) consists of a vaulted
room, 3.40 m. wide and 6.60 m. long, which is divided into twelve
compartments by one longitudinal and five baffle walls. The gases
change direction seven times, and pass over the longitudinal wall
six times, being struck six times by fine sprays of water. The six
atomizers for this purpose consume 1.5 liter of water per minute, of
which four-fifths is vaporized, while one-fifth flows off to the lower
water basin. By this means 10 to 15 per cent. of the total flue dust
is precipitated in the condensing chamber itself, and is removed from
time to time as mud through the lower openings, which are water-sealed.
The remainder of the volatilized water precipitates the flue dust
almost completely on the way to the stack, so that only a short column
of steam is visible at the mouth of the stack. The flue to the stack
passes for the most part underground through abandoned adits and
galleries, thus providing a variety of changes in cross-section and
in direction, and assisting materially the action of the condensing
chamber.
[Illustration: FIG. 47.—Section of Lead Refinery.]
[Illustration: FIG. 48.—Softening Furnace.]
As the charge of the shaft furnaces is poor in sulphur, no real matte
is produced, but only work lead and lead ashes (Bleischaum), which
contains 90 per cent. of lead, 1.6 per cent. sulphur, 0.4 per cent.
zinc, 0.85 per cent. Cu, 0.99 per cent. Fe, and 0.22 per cent. Sb. By
liquation and a reducing smelt in a reverberatory furnace, most of the
lead is obtained, along with a lead-copper matte, which is smelted for
copper matte and antimonial lead in the blast furnace.
[Illustration: FIG. 49.—Fume Condenser. (Section A B.)]
The copper matte, containing 18 per cent. Cu, 25 per cent. Fe, 30 per
cent. Pb and 18.4 per cent. S, is roasted dead in a reverberatory
furnace, is sintered, and melted to copper-bottoms in a small shaft
furnace. These copper-bottoms, which contain 60 per cent. copper and
25 per cent. lead, are subjected to liquation, and finally refined to
blister copper.
The zinc-desilvering plant, Fig. 47, consists of a reverberatory
softening furnace, two desilvering kettles of 14 tons capacity, a pan
for liquating the zinc crust, and a small kettle for receiving the lead
from the liquation process.
This pan has the advantage over the ordinary liquating kettle, that the
lead which drips off is immediately removed, before it can dissolve the
alloy; the silver content of the liquated lead is scarcely 0.05 per
cent., while the dry alloy contains 5 to 8 per cent.
[Illustration: FIG. 50.—Fume Condenser. (Section E F G H.)]
[Illustration: FIG. 51.—Fume Condenser. (Section C D.)]
The removal of the zinc is effected in a second reverberatory furnace.
Formerly the steam-method was used, but the rapid wear of the kettles,
and the excessive formation of oxides called for a change in the
process. The zinc-silver alloy is distilled in a crucible of 200 kg.
capacity, and is cupeled in an English cupel furnace. The details of
the reverberatory furnace are shown in Fig. 48.
The composition of the final products is shown by the following
analyses; Lead: Zn, 0.0021 per cent.; Fe, 0.0047 per cent.; Cu, 0.0005
per cent.; Sb, 0.0030 per cent.; Bi, 0.0007 per cent.; Ag, 0.0010 per
cent.; Pb, 99.998 per cent.; Silver, Ag, 99.720 per cent.; Cu, 0.121
per cent.; Fe, 0.005 per cent.; Pb, 0.018 per cent.; Au, 0.003 per
cent.
INDEX
Alloy, retorting the, in lead refining, 267
Alumina, experience with, 259
American Smelting and Refining Co., 4, 6, 26, 93, 113, 252, 295
at Murray, Utah, 287
Atmosphere, effect of on concrete, 242
Bag-house, cost of attending, 246
standard, 246
Bag-houses for saving fume, 244
Bartlett, Eyre O., 244
Bayston, W. B., 199
Bennett, James C., 66
Betts, Anson G., 270, 274
Between products, working up of, 39
Biernbaum, A., 41, 148, 160
Blast furnace of circular form, 253
Spanish lead, 307
Blast, volume and pressure of in lead smelting, 76
Blower, rotary, deficiency of, 251
Blowers for lead and copper smelting, 256
now more powerful for lead smelting use, 252
Blowers, rotary, method of testing volumetric efficiency of, 254
_vs._ blowing engines, 254
_vs._ blowing engines for lead smelting, 251
Blowing engines, when to use, 259
Bonne Terre lead deposits, 18
orebody, Missouri, 13, 14
Borchers, W., 114, 116, 127
Bormettes method, combination processes in, 222
Bradford, Mr., 55
Bretherton, S. E., 251, 258
Broken Hill Proprietary Block, 14, 59
Broken Hill practice, 51
Proprietary Co., 52, 113, 124, 145, 175, 178, 206
Bricking plant for flue dust and fine ores, 66-70
Briquetting costs, 62
methods of avoiding, 63, 64
process, operations, in 59
Bullion, analyses of in lead refining, 281
refined lead and slimes, analyses of, 282
Canadian Smelting Works, 275
Carlton Iron Co., 63
Carmichael, A. D. 56, 199
Carmichael-Bradford process, 175-185
brief estimate of, 209
claims of in patent, 199
recommendations of, 124
process, points concerning, 131
Cement walls, how to build, 241
Channing, J. Parke, 254
Charge-car in smelting, true function of, 94
feeding of in lead smelting, 77
mechanical character of in lead smelting, 78
Charges, effect of large in lead smelting, 77
Cherokee Lanyon Smelter Co., 104
Chimney bases, 237
Chisholm, Boyd & White Co., 64
Clark, Donald, 114, 144, 175
Cœur d’Alene mines, 5, 6, 7
Concrete flues and stacks, advantages and disadvantages of, 242
in metallurgical construction, 234
Connersville Blower Co., 252
Consolidated Kansas City Smelting and Refining Co., 285
Coke, percentage necessary to use in smelting, 259
Croll, H. V., 253
Cupellation in lead refining, 269
De Lamar Copper Refining Co., 297
Desilverization in lead refining, 265
Desloge practice contrasted with others, 46
Doeltz, F. O., 139
Dross, analyses of in lead refining, 279
Dupuis & Sons, 63
Dust chamber, arched form, 231
beehive form of, 232
design, 229
rectangular form, 230
concrete, 235-237
Dwight, Arthur S., 73, 81
spreader and curtain in furnaces, 91
East Helena and Pueblo smelting systems compared, 93
plant of the American Smelting and Refining Co., 302
system of smelting, 88-94
Edwards, Henry W., 234, 240, 242
Einstein silver mine, 14
Engine, blowing, proper field of, 257
blowing, and rotary blowers, 258
Eriksson, Hjalmar, 306
Federal Lead Co., 38
Mining and Smelting Co., 7
Feeders, cup and cone, for round furnaces, 81
Ferraris, Erminio, 311
Flat River mines, 18
Flue gases and moisture, effect of on concrete, 242
Flues, concrete, 234, 240, 242
Foundations for dynamos, 236
Fremantle Smelting Works, 145
Fume-smelting, cost of, 33
in the hearth, 32
Furnace operations at Desloge, Mo., 45
Furnaces at Desloge, Mo., 43
reverberatory, at Desloge, Mo., 42
Galena, experiments in roasting, 129
lime-roasting of, 14
new methods of desulphurizing, 116
roasting of by Savelsberg process, 122, 123
Gas, furnace, effect of on cement, 240
Gelatine, use of in electrolytic lead refining, 275
Germot, A., 224
process, 224
Globe plant of the American Smelting and Refining Co., 304
Smelting and Refining Co., 244
Greenway, T. J., 59
Guillemain, C., 133
Harvard, Francis T., 242
Hearth, covered-in, 36
Heat, effect of on cement, 242
Heberlein, Ferdinand, 113, 167, 199
Hixon, Hiram W., 256, 258
Harwood, E. J., 51
Hourwich, Dr. Isaac A., 27
Huntington-Heberlein process, 113, 144-147
consideration and estimate of, 203-209
credit due to, 126
process as distinguished from others, 118
economic results of, 155-159
Huntington-Heberlein explained by the inventors, 167-173
process at Friedrichshütte, 148
process, from the hygienic standpoint, 160
ideas of in patent specifications, 117
process, introduction of at Tarnowitz, Prussia, 41
and Savelsberg processes, essential difference between, 192
process, some disadvantages of, 165, 166
Huppertz, L., 121
Hutchings, W. Maynard, 108, 126, 170
Huntington, Thomas, 113, 167, 199
Iles, Malvern W., 96, 252
Ingalls, W. R., 3, 16, 27, 42, 177, 186, 193, 215, 224, 244, 287
Iron, behavior of in silver-lead smelting, 75
Jackson Revel mine, 14
Johnson, E. M., 104
R. D. O., 18
Jones, Richard, 244
Samuel T., 244
Laur, F., 224
Lead, analyses of refined, 281
bullion, electrolytic refining of base, 270
bullion, Parkes process of desilverizing and refining, 263
bullion, softening of, 263
concentrate Joplin district, valuation of, 25
and copper smelting, the Bormettes method of, 215-223
deposits, southeastern Missouri, 18
Joplin district, 8
marketing, 3
-ore roasting, consideration of new processes, 135-138
Lead ore, average prices for, 27
ore, cost of smelting, 32
-ore roasting, theoretical aspects of, 133
ores, Galena, Kan., 24
ores, method of valuing, 26
ores, southwestern Missouri, 24
Park City, Utah, 8
-poisoning in old and new processes, 162-165
refining, electrolytic, 274
soft, Missouri, 25
smelting at Desloge, Mo., 42
smelting at Monteponi, Sardinia, 311
smelting and refining, cost of, 96
smelting in the Scotch hearth, 31
smelting in Spain, 306
smelting at Tarnowitz, Prussia, 41
source of in Missouri, 13
in southeastern Missouri, 7, 10, 17
sulphide and calcium sulphate, metallurgical behavior of, 139-143
total production United States, 5
yield from Scotch hearths, 39
Leadville, Colo., mines, 8
Lewis, G. T., 244
Lime-roasting of galena, 126
Lotti, Alfredo, 215
Messiter, Edwin H., 229, 240
Middleton, K. W. M., 31
Mine La Motte, 14
Minerals, briquetting of, 63
Mining methods in Missouri, 19-23
Missouri Smelting Co., 197
Mould, H. S., Co., 64
Murray smelter, Utah, 291
National plant of the American Smelting and Refining Co., 299
New Jersey Zinc Co., 246
Nutting, Mr., 256
Ore and Fuel Co., 63
different behavior of coarse and fine in lead smelting, 79
treatment in detail by the Huntington-Heberlein process, 150-155
Parkes process, cost of refining by, 99
Percy, Dr., 244
Perth Amboy plant of the American Smelting and Refining Co., 296
Petraeus, C. V., 24
Pfort curtain for furnaces, 82
Picher Lead Co., 197
Piddington, F. L., 263
Potter, Prof. W. B., 15
Pueblo lead smelter, 294
Smelting and Refining Co., 84
Pufahl, O., 38, 291, 294, 296, 299, 302, 304
Pyritic smelting without fuel practically impossible, 256
Raht, August, 251, 254
Refining, monthly cost of per ton of bullion treated, 100
Roasters, hand, and mechanical furnaces, average monthly cost of, 98
Roberts-Austen, W. C., 139
Salts, effect of crystallization of contained on concrete, 243
Santa Fe Gold and Copper Mining Co., 255
Savelsberg, Adolf, 122
Savelsberg process, 186-192
process, claims of in patent, 201
process contrasted with Huntington-Heberlein, 209
process, difference between and Huntington-Heberlein, 197
Savelsberg process the simplest, 132
Scotch-hearth method, permanency of, 195
Scotch hearths, 34
Schneider, A. F., 81
Seattle Smelting and Refining Works, 273
Silver-lead blast furnaces, mechanical feeding of, 81
blast furnace, proper conditions, 73
smelting, details of practice, 73
smelting, modern, 73
Slag-smelting costs, 34
Slime analysis at Broken Hill, 51
Slimes, analyses of in lead refining, 281
desulphurization of by heap roasting, 51
treatment of at Broken Hill, 53-55
Smelter, new, at El Paso, Texas, 285
Smelters’ pay, 32
Smelting, average cost of per ton, 98
Smelting Co. of Australia, 263
costs, 48
detailed costs of, 101, 102
of galena ore, 38
preparation of fine material for, 59
Solution, washing from slime, 277
Sticht, Mr., 256
St. Joseph Lead Co., 16
St. Louis Smelting and Refining Co., 81
Sulphide Corporation, 145
Sulphur dioxide, effect of on cement, 240
Sulphuric acid, making of at Broken Hill, 174
Tasmanian Smelting Co., 145
Tennessee Copper Co., 255
Terhune, R. H., furnace gratings, 84
Thacher, Arthur, 14
Ulke, Titus, 270
United Smelting and Refining Co., 88
States Zinc Co., 295
Vezin, H. A., 252
Walls, retaining, 237
Walter, E. W., 260
Waring, W. Geo., 24
Welch, Max J., 229
Wetherill, Samuel, 244
Wheeler, H. A., 10
Zinc, amount required in lead refining, 265, 266
crusts, treatment of in lead refining, 267
oxide in slags, 108
retort residues, analysis of materials smelted and
bullion produced, 106
retort residues, smelting, 104
FOOTNOTES:
[1] During 1905, antimonial lead commanded a premium of about 1c. per
lb. above desilverized, owing to the high price for antimony.
[2] The figures for 1903 and 1904 have been added in the revision of
this article for this book. The production of lead in the United States
in 1903 was 276,694 tons; in 1904, it was 302,204 tons.
[3] Ounces of silver to the ton of lead.
[4] These figures are doubtful; they are probably too high. (See table
on p. 5).
[5] The production of zinc ore in this district has now been commenced.
[6] The manuscript of this article was dated Oct. 5, 1905.
[7] Translated from _Zeit. f. Berg.-Hütten-und Salinenwesen_, LIII
(1905, p. 450).
[8] This paper is published in pp. 148-166 of this book.
[9] Abstract from _Transactions_ of the Australasian Institute of
Mining Engineers, Vol. IX, Part 1.
[10] In the course of subsequent discussion Mr. Horwood stated that the
losses in roasting were 12½ per cent. in lead and probably about 5 per
cent. in silver. As compared to roasting in Ropp furnaces the loss in
lead was 5 to 6 per cent. greater, but the difference of loss in silver
was, he thought, not appreciable. Mr. Hibbard said that the Central
mine had obtained satisfactory results with masonry kilns.—EDITOR.
[11] Abstract of portion of a paper presented at the Mexican meeting
of the American Institute of Mining Engineers, under the title “The
Mechanical Feeding of Silver-Lead Blast Furnaces.” _Transactions_, Vol.
XXXII, pp. 353-395.
[12] Abstract of a paper (“The Mechanical Feeding of Silver-Lead Blast
Furnaces”) presented at the Mexican meeting of the American Institute
of Mining Engineers and published in the _Transactions_, Vol. XXXII.
For the first portion of this paper see the preceding article.
[13] Abstract of a paper in _Western Chemist and Metallurgist_, I, VII,
Aug., 1905.
[14] Much better work is being done at present, smelting the Western
zinc ores, and the residue contains about one-third of the above
figure, or 7.5 per cent. of zinc oxide. The high per cent. of ZnO left
in residue was mainly due to poor roasting.
[15] There was also considerable coke used of an inferior grade, made
from Kansas coal.
[16] Part of the ZnO in roasted matte came from being roasted in the
same furnace the zinc ore had been roasted in.
[17] There was less residue on the charges during this month, which
accounts for the larger tonnage with a lower blast.
[18] Translation of a paper read before the Naturwissenschaftlicher
Verein at Aachen, and published in _Metallurgie_, 1905, II, i, 1-6.
[19] 35 to 40 cm. = 13.78 to 15.75 in. = 8 to 9.12 oz. per sq. in.
[20] _Engineering and Mining Journal_, 1904, LXXVIII, p. 630; article
by Donald Clark; reprinted in this work, p. 144.
[21] Owner of the patents.—EDITOR.
[22] Abstract of a paper in _Metallurgie_, II, 18, Sept. 22, 1905, p.
433.
[23] This method is described further on in this book.
[24] Translated from _Metallurgie_, Vol. II, No. 19.
[25] British patent, No. 17,580, Jan. 30, 1902, “Improved process for
desulphurizing sulphide ores.”
[26] W. C. Roberts-Austen, “An Introduction to the Study of
Metallurgy,” London, 1902.
[27] A. Lodin, _Comptes rendus_, 1895, CXX, 1164-1167; _Berg. u.
Hüttenm. Ztg._, 1903, p. 63.
[28] _Comptes rendus_, loc. cit.
[29] Translated from the _Zeitschrift für das Berg.-Hütten-und
Salinenwesen im. preuss. Staate_, 1905, LIII, ii, pp. 219-230.
[30] Translated from the _Zeitschrift für das Berg.-Hütten-und
Salinenwesen im. preuss. Staate_, 1905, LIII, ii, pp. 219-230.
[31] The manufacture of sulphuric acid from these gases has now been
undertaken in Silesia on a working scale.—EDITOR.
[32] A paper presented before the American Institute of Mining
Engineers, July, 1906.
[33] _Engineering and Mining Journal_, Sept. 2, 1905.
[34] This term is inexact, because the hearths employed in the United
States are not strictly “Scotch hearths,” but they are commonly known
as such, wherefore my use of the term.
[35] Percentages of lead in Missouri practice are based on the wet
assay; among the silver-lead smelters of the West the fire assay is
still generally employed.
[36] This improvement did not originate at either Alton or
Collinsville. It had previously been in use at the works of the
Missouri Smelting Company at Cheltenham, St. Louis, but the idea
originated from the practice of the Picher Lead Company, of Joplin, Mo.
[37] This refers especially to the Savelsberg process.
[38] A. D. Carmichael, U. S. patent No. 705,904, July 29, 1902.
[39] _Metallurgie_, 1905, II, i, 1-6; _Engineering and Mining Journal_,
Sept. 2, 1905.
[40] _Metallurgie_, 1905, II, 19; _Engineering and Mining Journal_,
Jan. 27, 1906.
[41] _Metallurgie_, 1905, Sept. 22, 1905; _Engineering and Mining
Journal_, March 10, 1906.
[42] _Engineering and Mining Journal_, Oct. 21, 1905.
[43] Translated by W. R. Ingalls.
[44] As originally published the title of this article was
“Lead-Smelting without Fuel.” In this connection reference may well be
made to Hannay’s experiments and theories, _Transactions_ Institution
of Mining and Metallurgy, II, 188, and Huntington’s discussion,
_ibid._, p. 217.
[45] Excerpt from a paper, “Concrete in Mining and Metallurgical
Engineering,” _Transactions_ American Institute of Mining Engineers,
XXXV (1905), p. 60.
[46] A Discussion of the Paper by Henry W. Edwards, on “Concrete in
Mining and Metallurgical Engineering,” _Transactions_ of the American
Institute of Mining Engineers, XXXV.
[47] _Engineering News_, Nov 30, 1899, and U. S. Patent No. 665,250,
Jan. 1 1901.
[48] A discussion of the paper of Henry W. Edwards, on “Concrete in
Mining and Metallurgical Engineering,” _Transactions_ of the American
Institute of Mining Engineers, XXXV.
[49] Abstract from the _Journal_ of the Chemical, Metallurgical and
Mining Society of South Africa, May, 1903.
[50] Abstract of a paper in _Transactions_ American Institute of Mining
Engineers, XXXIV (1904), p. 175.
[51] Silver not given. This was the case, also, with the gold in the
bullion. The slimes contained 0.131 per cent. of gold, or 39.1 oz. per
ton.
[52] A constituent company of the American Smelting and Refining
Company.
[53] Translated from _Zeit. f. Berg.-Hütten.-und Salinenwesen im
preuss. Staate_, 1905, LIII, p. 433.
[54] Abstract from a paper in _Zeit. f. Berg.-Hütten-und Salinenwesen
im preuss. Staate_, 1905, LIII, p. 439.
[55] Translated from _Zeit. f. Berg.-Hütten.-und Salinenwesen im
preuss. Staate_, 1905, LIII, 490.
[56] Abstract from a paper in _Zeit. f. Berg.-Hütten-und Salinenwesen
im preuss. Staate_, 1905, p. 400.
[57] Abstract from a paper in _Zeit. f. Berg.-Hütten.-und Salinenwesen
im preuss. Staate_, 1905, p. 400.
[58] Abstract from an article in _Zeit. f. Berg.-Hütten.-und
Salinenwesen im preuss. Staate_, 1905, LIII, p. 444.
[59] Translated from _Oest. Zeit. f. Berg.-und Hüttenwesen_, 1905, p.
455.
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