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By E. H. Erwin Gartner

The Rhenish lignite deposit in the Nordrhein-Westphalia province of the German Federal Republic covers an area on the left bank of the Rhine River of about 970 sq miles (Fig. 1). Here, in the Miocene era, some 20 million years ago a lignite seam was formed and later broken, through tectonic movements into several blocks, some thrown down into depths of over 1,500 ft. This deposit consists of a single seam, in places up to 300 ft thick, but averaging about 150 ft. In a few areas, it is split into as many as five seams, separated by barren strata. It contains about 60 billion tons (2000 lb/short ton) of lignite which, however, being relatively young geologically, has, because of approximately 60% moisture, 2 to 8% ash, and low sulfur, a net calorific value of 3,340 Btu. That calorific value increases, however, with rising thickness of the cap rock. About one-third of the yearly production of 90 to 100 million tons of raw lignites is converted into briquets. Raw lignite is pulverized and dried to about 18% water content and then pressed without binder into briquets under a pressure of about 14,000 psi. The briquets are being used primarily for residential heating; their sales, however, have been declining sharply since 1963. On the other hand, the volume of raw lignite used in steam power plants has been rising steadily and amounts at present (1968) to about two-thirds of the total production. Huge steam power plants, such as Frimmersdorf, shown in Fig. 2 (which, with an installed capacity of 2,600 MW [megawatt], represents the world's largest steam power plant) with a generating capacity of up to 300 MW per unit and a total installed capacity of over 6,000 MW, contribute about 26% of the total electric power production of the German Federal Republic.

Jan 1, 1969

By J. Moore Samuel

THE prospective change in ore receipts at the Copper Queen reduction works necessitated a careful study of conditions to determine the most economical method of smelting. The first step, calculating the known, probable, and possible results, showed the economy of a unimetallurgical flow sheet f which the three pyrometallurgical stages would be roasting, reverberatory smelting, and converting. The past and future ore receipts (not including converter flux) are as follows: AVERAGE MONTHLY FUTURE MONTHLY RECEIPTS, 1918 RECEIPTS TONS PER PER CENT. OF TONS PER PER CENT. OF MONTH TOTAL RECEIPTS MONTH TOTAL RECEIPTS Ore 84,553 82.3 43,000 44.0 Gravity concentrates 16,375 15.9 25,500 26.0 Flotation concentrates 1,883 1.8 29,400 30.0 Total 102,811 100.0 97,900 100.0 Of the 102,811 tons per month received in 1918, only 28,945, or 28 per cent., was treated in the roasters and an average f 14.1 roasters out of 20 were in operation; the remaining ore was treated in the blast furnaces. The change to "all reverb." practice, therefore, called for considerable additional roaster capacity, which could be obtained either by building additional units, or by remodelling the present units so as to give increased unit capacity. In view of the progress in roasting practice since the construction of the Copper Queen roaster plant and the heavy investment in that roaster plant, the second plan seemed to be the more desirable. The Copper Queen roaster plant consisted of twenty six-hearth, McDougall roasters, 18 ft. in diameter, equipped with cast-iron arms that originally were water cooled but now are air cooled. They have no top dry hearth and the feed is discharged through the top arch, near its periphery, by plunger feeders from two bins at opposite sides of the roaster.

Jan 7, 1921

By AIME AIME

THAT the copper industry is in serious straits is admitted. So are the lead and zinc industries, and both lead and zinc are tariff protected. Conditions in the Western lead, zinc and silver mining districts are desperate. These are illustrated by the present disastrously low prices of these metals compared with the levels of two and three years ago and by the comparative rates of metal production and the present activity of the western mining camps. The January production of copper was 32.2 per cent of the 1929 average rate, of steel was 32.3 per cent of the 1929 average, while January production of pig iron was only 27.6 per cent of the monthly average of 1929. The copper industry is only one of many industries that are suffering in this tragic depression. It has been exceedingly prosperous and will be prosperous again when business recovers. Its difficulties are not unique and it does not deserve the tariff equivalent of a dole. The burden of subsidizing through an effective tax or tariff the minority group of high-cost mines would be a heavy one on the consuming public of the United States. On the basis of 1929 consumption figures, an effective 5c. tariff would result in an increased copper bill to the country of over $109,000,000, of which the electric manufacturers would contribute $26,000,000, the telephone and telegraph companies $16,000,000, and the automobile manufacturers $14,000,000. And of this total bill, only $33,000,000 would go to those who are urging a tariff and the balance of $76,000,000 would be paid to those who do not admit either its necessity or usefulness under present conditions.

Jan 1, 1932

By A. A. Smith, J. S. Smart

THe elements silver, cadmium, tin, antimony and tellurium either are found as impurities in commercial coppers or are intentionally added to produce coppers for special uses. When present in small quantities, each of these metals exerts a significant effect on the properties of copper, but despite considerable experience and investigation much remains to be learned about their specific behavior. The presence of extraneous elements, particularly oxygen, complicates the study of individual additions of impurities in the range o to 0.05 per cent, and the use of high-purity copper has been found extremely helpful in circumventing these difficulties. It has also been found that individual behavior can be satisfactorilv determined from measurements of two commercially important properties, conductivity and softening temperature, according to methods that have been described previously.1,2 Data of this type, obtained in the course of a general investigation of the effects of impurities on the properties of both oxygen-bearing 2nd oxygen-free copper are presented herewith. All samples except those containing cadmium were prepared as 3/8-in. continuously cast oxygen-free rod from pure copper and weighed amounts of a similarly synthesized master alloy. Composition was checked by chemical methods, which in every case verified the calculated analyses. Test wires were cold-drawn one B, and S. number per pass on a string-up machine using four 30-min. 600°C. anneals at 0.3125, 0.257, 0.204 and 0.162 in. and a final cold reduction of 75 per cent to 0.081 in. Some oxygen-bearing samples were produced directly from the 3/8-in. oxygen-free rods by diffusion from a surface scale at 850°C. and drawn to size as listed above, but most of the alloys were remelted in pure graphite and cast through air into a I-in. round mold. The slugs were then hot-rolled to 5/16 in., drawn to 0.162 in., annealed as listed above, and drawn to 0.081 inch. Conductivity wires were annealed for I hr. at 100°C. intervals between 300' and 800°F., using a hydrogen atmosphere for oxygen-free samples and purified nitrogen for the oxygen-bearing type. All samples were quenched from the annealing temperature in 10 per cent H²S04 and conductivity was measured on a Hoopes bridge that had been checked against a Kelvin bridge. The softening temperatures represent the half-hard stage of the annealing curve as determined by a standard procedure.' Effect of Silver Numerous excellent studies of the argentiferous coppers are found in the literature, and these have been adequately

Jan 1, 1943

By A. A. Smith, J. S. Smart

THe elements silver, cadmium, tin, antimony and tellurium either are found as impurities in commercial coppers or are intentionally added to produce coppers for special uses. When present in small quantities, each of these metals exerts a significant effect on the properties of copper, but despite considerable experience and investigation much remains to be learned about their specific behavior. The presence of extraneous elements, particularly oxygen, complicates the study of individual additions of impurities in the range o to 0.05 per cent, and the use of high-purity copper has been found extremely helpful in circumventing these difficulties. It has also been found that individual behavior can be satisfactorilv determined from measurements of two commercially important properties, conductivity and softening temperature, according to methods that have been described previously.1,2 Data of this type, obtained in the course of a general investigation of the effects of impurities on the properties of both oxygen-bearing 2nd oxygen-free copper are presented herewith. All samples except those containing cadmium were prepared as 3/8-in. continuously cast oxygen-free rod from pure copper and weighed amounts of a similarly synthesized master alloy. Composition was checked by chemical methods, which in every case verified the calculated analyses. Test wires were cold-drawn one B, and S. number per pass on a string-up machine using four 30-min. 600°C. anneals at 0.3125, 0.257, 0.204 and 0.162 in. and a final cold reduction of 75 per cent to 0.081 in. Some oxygen-bearing samples were produced directly from the 3/8-in. oxygen-free rods by diffusion from a surface scale at 850°C. and drawn to size as listed above, but most of the alloys were remelted in pure graphite and cast through air into a I-in. round mold. The slugs were then hot-rolled to 5/16 in., drawn to 0.162 in., annealed as listed above, and drawn to 0.081 inch. Conductivity wires were annealed for I hr. at 100°C. intervals between 300' and 800°F., using a hydrogen atmosphere for oxygen-free samples and purified nitrogen for the oxygen-bearing type. All samples were quenched from the annealing temperature in 10 per cent H²S04 and conductivity was measured on a Hoopes bridge that had been checked against a Kelvin bridge. The softening temperatures represent the half-hard stage of the annealing curve as determined by a standard procedure.' Effect of Silver Numerous excellent studies of the argentiferous coppers are found in the literature, and these have been adequately

Jan 1, 1943

By Alden B. Greninger, Victor G. Mooradian

Microscopic examinations of quenched (beta) 60:40 copper-zinc and quenched (beta) 74:26 copper-tin alloys have frequently disclosed long, thin parallel bands or markings within the polyhedral beta grains. Apparently, it has been accepted that these markings are mechanical twin bands, although there is no published account of experimental data that might tend to substantiate this belief for the copper-zinc system, and only meager evidence has been submitted for the copper-tin system. Johnsonl0 commented upon the markings he observed after deforming a quenched brass of 55.6 per cent copper and 44.4 per cent zinc. Clark's papers4 on the heat-treatment of Muntz metal contain many photomicrographs of markings in beta brass; these she referred to as Neumann bands (inasmuch as repolishing and re-etching failed to remove them) and stated that markings present in Muntz metal as quenched were the result of shock imparted during the polishing operation. Van Wert18 first suggested that these parallel markings in quenched beta brass (60:40) might be alpha precipitated along certain crystallographic planes of the matrix lattice. He showed that upon reheating the specimens to temperatures below 800" C., obvious precipitation of alpha within the large beta grains first occurred at the markings. In a discussion of this paper, Phillips stated (pp. 505-507) that in his opinion the markings observed by Van Wert were mechanical twins resulting from contraction strains produced during the quenching operation. In a study of segregate structures in beta copper-tin, Smith17 observed that in the quenched beta phase (containing 74.92 per cent Cu) parallel . bands or markings very similar to those in metastable beta brass could be

Jan 1, 1938

By Alden B. Greninger, Victor G. Mooradian

Microscopic examinations of quenched (beta) 60:40 copper-zinc and quenched (beta) 74:26 copper-tin alloys have frequently disclosed long, thin parallel bands or markings within the polyhedral beta grains. Apparently, it has been accepted that these markings are mechanical twin bands, although there is no published account of experimental data that might tend to substantiate this belief for the copper-zinc system, and only meager evidence has been submitted for the copper-tin system. Johnsonl0 commented upon the markings he observed after deforming a quenched brass of 55.6 per cent copper and 44.4 per cent zinc. Clark's papers4 on the heat-treatment of Muntz metal contain many photomicrographs of markings in beta brass; these she referred to as Neumann bands (inasmuch as repolishing and re-etching failed to remove them) and stated that markings present in Muntz metal as quenched were the result of shock imparted during the polishing operation. Van Wert18 first suggested that these parallel markings in quenched beta brass (60:40) might be alpha precipitated along certain crystallographic planes of the matrix lattice. He showed that upon reheating the specimens to temperatures below 800" C., obvious precipitation of alpha within the large beta grains first occurred at the markings. In a discussion of this paper, Phillips stated (pp. 505-507) that in his opinion the markings observed by Van Wert were mechanical twins resulting from contraction strains produced during the quenching operation. In a study of segregate structures in beta copper-tin, Smith17 observed that in the quenched beta phase (containing 74.92 per cent Cu) parallel . bands or markings very similar to those in metastable beta brass could be

Jan 1, 1938

By F. E. Lathe, L. Hodnett

This paper summarizes extensive data supplied by 40 copper converter plants in 18 countries, and includes a partial analysis and comments on the effect of converter slag composition and temperature on the fomzation and elimination of magnetite and the life of basic and neutral refractories. APRELIMINARY draft of this paper, including data from 26 companies which had submitted replies to a questionnaire, was presented at the Annual Meeting of the AIME held in New York in February, 1956. Publication was delayed in the expectation that additional replies would be received, and justification for the subsequent delay may be found in the fact that the present revised paper contains data from 40 converter plants situated in 18 different countries. The suggestion that a questionnaire be sent to the major copper and copper-nickel plants outside the U.S.S.R. and associated countries arose at the Los Angeles Annual Meeting in February, 1953, where several papers dealing with converter construction and operation were presented. In private discussions which followed, a group of metallurgists and refractories engineers raised the question why practices varied so much at different plants. Why did magnesite brick give superior service to chrome-magnesite in some plants but chrome-magnesite prove superior in others? Would metallurgical studies point the way to the development of still better refractories? Could magnetite formation be prevented in converter slags and reverber-atory furnace mattes? If this was feasible, why was a heavy circulating load of magnetite tolerated, when, in all probability, it increased copper losses in the reverberatory slags and, in some cases, made it necessary to shut down the furnaces periodically to clean out the accumulated magnetite? Was the sole purpose of coating brick with magnetite that of

Jan 1, 1959

By H. H. Richardson, E. H. Dix

Of all the alloying elements used in commercial aluminum alloys, copper stands out as by far the most important, and it is perhaps for this reason that the constitution of the aluminum-copper system has received the attention of a number of experimenters covering a period of approximately 20 years. It would be expected that during this time the equilibrium relations would have been thoroughly and exactly worked out. In a general way this is true, but there is a surprising lack of agreement in the details of the high aluminum end of the diagram, particularly in regard to the solid solubility of copper in aluminum, which has been given variously as 2 per cent. to 5 per cent. at the eutectic temperature. The importance of this is much greater than might at first appear to one unfamiliar with modern commercial aluminum alloys. This is true because, with one exception, the susceptibility of the strong alloys to heat treatment depends, in part at least, upon the solid solubility relations of copper in aluminum. In 1919, Merica, Waltenberg, and Freeman1 published the first detailed information on these solubility relations. They found that the solubility of copper decreases with falling temperature from about 4 per cent. at 525" C. to about 1 per cent. at 300" C. and is apparently still diminishing at lower temperatures. In 1921 there was published a brief account of the results obtained in the National Physical Laboratory, Teddington, England.2 This laboratory, however, reported that 5 per cent. of copper is soluble in aluminum at 540" C., and that this amount decreases to about 3 per cent. at 20" C. In the same year Ohtani and Hemmi3 reported a solubility of 4.8 per cent. copper at 520" C., 2.6 per cent. at 460" C. and 1.5 per cent. at 420" C.

Jan 1, 1926

By Wheeler North

THE United Eastern Mining Co.'s property is located in the Oatman, Gold Roads mining district of Mohave County, Ariz. The mine and mill are some 26 mi. (41.8 km.) southwest of Kingman, the nearest railway connection. The ore consists of a mixture of calcite and quartz with some andesite. The values, which are mainly gold, alloyed with approximately 34 per cent., by weight, of silver, are extremely fine so that fine grinding is a most important factor in extraction. The milling process, briefly, consists of single-stage, coarse-crushing, two-stage ball milling in cyanide solution, followed by combined mechanical and air agitation, and by straight counter-current decantation. The Merrill zinc-dust precipitation method is used to recover the values, and is preceded by the Crowe vacuum treatment of the solution. The most notable features of the plant are the absence of filters, and the use of short peb mills loaded with steel balls, for fine grinding. As originally designed, the capacity of the mill was to be 200 tons per day, although the grinding end was to be capable of handling double that amount. By adding extra agitators and thickeners and by some alterations to the grinding end, the daily capacity has been raised to 285 tons. This, it is hoped, may be increased to 300 tons during the warm months of the present year. TONNAGE MILLED PER DAY 1917 average 234 Maximum month 272 Minimum month 146 1918 average 253 Maximum month 276 Minimum month 217 As this article primarily deals with plant operations, those interested in the mechanical description are referred to the article by Mr. Otto Wartenweiler,1 the designing engineer, which gives the details and costs of design and construction. The following changes have been made since that article was published:

Jan 8, 1919

By Joseph Bernhardt

THE Cornwall Ore Mines, Division of the Bethlehem Steel Co., at Cornwall, Lebanon County, consists of two separate magnetite ore bodies, approximately one mile apart. The one ore body was an outcrop in which open-pit mining operations were begun in 1742 and continued without interruption. This mine is also served by a shaft for mining that portion of the ore body not recoverable by open-pit methods. The other property had no outcropping and is entirely an underground operation. This mine, cap-tioned No. 4 mine, is serviced by two shafts 220 ft apart. The No. 4 shaft is for hoisting rock and ore, and the No. 5 shaft is for handling men and ma- terials. The shafts are on a catenary curve, closely following the dip of the ore body. Slope of the shafts at the collar is 36", flattening to 26" at the bottom. The footwall at Cornwall is a traprock which varies structurally from solid to highly fractured. Grizzly stations at the top of each chute raise were reinforced by setting timber sills on concrete. Occasionally it was necessary to concrete the entire chute raise. Concreting was done by using 35-yd electric mixers set up in the breaking subs. Sand, stone, and cement were hoisted up through the chute raises to the breaking subs. Late in 1937 it was decided to eliminate chute raises and breaking subs in future development. Haulage levels are 50 ft apart vertically. When ore is below a 40 " dip, slushing drifts are driven normal to the haulages and on 40-ft centers. In a few areas, ore dips greater than 40" and slushing drifts are then placed parallel to the haulage drifts. Finger raises are opposite each other and on 20-ft centers making 20x20-ft draw areas (fig. 1). For loading

Jan 1, 1951

By Joseph Bernhardt

THE Cornwall Ore Mines, Division of the Bethlehem Steel Co., at Cornwall, Lebanon County, consists of two separate magnetite ore bodies, approximately one mile apart. The one ore body was an outcrop in which open-pit mining operations were begun in 1742 and continued without interruption. This mine is also served by a shaft for mining that portion of the ore body not recoverable by open-pit methods. The other property had no outcropping and is entirely an underground operation. This mine, cap-tioned No. 4 mine, is serviced by two shafts 220 ft apart. The No. 4 shaft is for hoisting rock and ore, and the No. 5 shaft is for handling men and ma- terials. The shafts are on a catenary curve, closely following the dip of the ore body. Slope of the shafts at the collar is 36", flattening to 26" at the bottom. The footwall at Cornwall is a traprock which varies structurally from solid to highly fractured. Grizzly stations at the top of each chute raise were reinforced by setting timber sills on concrete. Occasionally it was necessary to concrete the entire chute raise. Concreting was done by using 35-yd electric mixers set up in the breaking subs. Sand, stone, and cement were hoisted up through the chute raises to the breaking subs. Late in 1937 it was decided to eliminate chute raises and breaking subs in future development. Haulage levels are 50 ft apart vertically. When ore is below a 40 " dip, slushing drifts are driven normal to the haulages and on 40-ft centers. In a few areas, ore dips greater than 40" and slushing drifts are then placed parallel to the haulage drifts. Finger raises are opposite each other and on 20-ft centers making 20x20-ft draw areas (fig. 1). For loading

Jan 1, 1951

By A. L. Blomfield, M. J. Trott

ThE Golden Cycle Mining and Reduction Co. operates its custom mill at Colorado Springs on Cripple Creek ores exclusively. These ores are straight sulfo-tellurides, with practically no base metals such as Pb, CU, Zn, As, or Sb. Upon entering the mill, the ore all passes through the sampler to storage bins, where beds of 5000 tons each are made to insure uniformity in roasting. Classification of ORes For efficiency, it has been found advisable to separate the ores into two classes according to their lime content. Other constituents have their effect, such as alumina and magnesia, but, speaking generally, these are roughly proportional to the lime, which, as CaO, is used as the indicator. Class A contains not more than 2 per cent. CaO and comprises about 66 per cent. of the ore treated. Class B contains over 2 per cent. CaO, and comprises 34 per cent. of the ore treated. Typical analyses of both classes are as follows: Class A a B Insol.......................... 86.70 75.90 Al2O3........................:.. 2.30 3.40 Fe.............................. 3.26 3.48 CaO............................. 1.57 5.12 S............................... 1.79 1.80 MgO........................... 0.21 1.08 Ignition loss..................... 3.20 6.50 99.03 97.28 The beds of each class of ore are made as large as possible with the available ore as it comes in, to avoid changing from one bed to the other any oftener than necessary. Each roaster will handle 50 per cent. more of class A ore than of class B, with practically the same results. This necessitates changing the rate of feed for each new bed and tends toward a poor roast and loss of tonnage at each change.

Jan 1, 1919

By Harold F. Unger, Byron S. Snowden, William H. Engelmann

The effects of using surfactant solutions while diamond drilling in amygdaloidal basalt of the Upper Michigan copper mining district were investigated. Nonionic, anionic, and cationic surfactant solutions, all biodegradable, were tested with size AX surface-set bits mounted on a drill instrumented for continuous recording of thrust, rotational speed, and torque levels. Water flow and penetrations rates were also recorded. A drilling study was carried out, testing two rotational speed levels, two diamond bit grade levels, and four flushing agents (including water as a control). Comparison of all low-speed/high-thrust groups of tests shows that the non-ionic surfactant produced a 35% overall average increase in penetration rate over water, the anionic a 29% increase, and the cationic a 28% increase. These increases in, penetration rates resulted in a 22 to 26% reduction in cost over flushing with water alone. With the high-speed/low-thrust combination, the surfactants produced less benefits in penetration rate increases over water (14% increase for nonionic, 14% decrease for anionic, and 2% decrease for cationic type). Particle size analyses were run on drill cuttings produced by the three surfactant solutions and by water. All surfactant solutions showed a 20 to 50% increase in solids content over that of water alone. Since flushing rates were constant, these increases in solids content agree well with the penetration rate changes. The diamonds were recovered from each bit by matrix dissolution, and bit wear was measured in terms of total loss of weight. Flushing with surfactant solutions showed a savings in bit wear in most cases. The drill team noted also that surfactant flushing improved recovery of drill core by allowing easier removal of the core from the barrel.

Jan 1, 1976

By A. L. Blomfield

THE Golden Cycle Mining and Reduction Co. operates its custom mill at Colorado Springs on Cripple Creek ores exclusively. These ores are straight sulpho-tellurides, with practically no base metals such as Pb, Cu, Zn, As, or Sb. Upon entering the mill, the ore all passes through the sampler to storage bins, where beds of 5000 tons each are made to insure uniformity in roasting. CLASSIFICATION OF ORES For efficiency, it has been found advisable to separate the ores into two classes according to their line content. Other constituents have their effect, such as alumina and magnesia, but, speaking generally, these are roughly proportional to the line, which, as CaO, is used as the indicator. Class A contains not more than 2 per cent. CaO and comprises about 66 per cent. of the ore treated. Class B contains over 2 per cent. CaO, and comprises 34 per cent. of the ore treated. Typical analyses of both classes are as follows: Class A Class B Insol 56.70 75.90 Al203 2.30 3.40 Fe 3.26 3.48 CaO 1.57 5.12 S 1.79 1.80 Mg0 0.21 1.08 Ignition loss 3.20 6.50 99.03 97.28 The beds of each class of ore are made as large as possible with the available ore as it comes in, to avoid changing from one bed to the other any oftener than necessary. Each roaster will handle 50 percent. more of class A ore than of class B, with practically the same results. This necessitates changing the rate of feed for each new bed and tends toward a poor roast and loss of tonnage at each change.

Jan 8, 1918

By W. Geo. Waring

Introduction THE winning of zinc and lead ores from the comparatively shallow deposits of the Joplin district, presents, few such problems for the mining engineer as are encountered in deep ore mining and in the handling of complex ores. However, for those who are interested in the origin and nature of ore deposits in general, there is much here to investigate, and upon studies in this direction will depend in no small measure the future of an industry that has produced during the year 1916, in southwest Missouri, lead and zinc ores to the value of $26,319,383; adding $46,022,164 to the metallic wealth of the nation. Including the adjacent mines in Ottawa County, Oklahoma, and Cherokee County, Kansas, which are, really continuations of the Joplin district, the total value of ores produced reached the sums of $34,089,503, and that of the metals, $58,927,941.1 The ores produced in the Joplin district are remarkable on account of their almost uniformly high" metal content in both lead and zinc. Their superiority in this regard is unparalleled, and is doubtless clue to certain geological conditions that are now possibly explicable, provided that the artesian circulation = ascension hypothesis of the origin of the ore deposits is proven beyond question to be the only tenable explanation of the facts. These, as well as other valuable zinc-ore deposits, are bound to he exhausted in time, for ordinary mining purposes, acid it concerns the mining engineer that the conditions under which ores of this character were actually deposited in the form in which they occur, may, if possible, be ascertained, in order that he may learn where to look for similar deposits in other regions, or even how to reproduce artificially such condi-

Jan 9, 1917

By W. Geo. Waring

The winning of zinc and lead ores from the comparatively shallow deposits of the Joplin district presents few such problems for the mining engineer as are encountered in deep ore mining and in the handling of complex ores. However, for those who are interested in the origin and nature of ore deposits in general, there is much here to investigate, and upon studies in this direction will depend, in no small measure, the future of an industry that has produced during the year 1916, in southwest Missouri, lead and zinc ores to the value of $26,319,383; adding $46,022,164 to the metallic wealth of the nation. Including the adjacent mines in Ottawa County, Oklahoma, and Cherokee County, Kansas, which are really continuations of the Joplin district, the total value of ores produced reached the sum of $34,089,503, and that of the metals, $58,927,941.l The ores produced in the Joplin district are remarkable on account of their almost uniformly high metal content in both lead and zinc. Their superiority in this regard is unparalleled, and is doubtless due to certain geological conditions that are now possibly explicable, provided that the artesian circulation = ascension hypothesis of the origin of the ore deposits is proved beyond question to be the only tenable explanation of the facts. These, as well as other valuable zinc-ore deposits, are bound to be exhausted in time, for ordinary mining purposes, and it concerns the mining engineer that the conditions under which ores of this character were actually deposited in the form in which they occur, may, if possible, be ascertained, in order that he may learn where to look for similar deposits in other regions, or even how to reproduce artificially such condi-

Jan 1, 1918

By H. E. Montgomery

H. L. Burghoff.*—I have a question, Dr. Smith. A number of the alloys you mentioned contain lead, which, of course, is present as discrcte particles. At what level of concentration does lead begin to play an important part in determining these elevated-temperature characteristics ? A. J. Smith.—That is a difficult question to answer because of the complicating features of the alloys that arc lead free. I do not know of any work done on the subject. But for comparison, 85-5-5-5 has about the same characteristics at, say 450°F. as 88-10-2, which is lead free or contains at most 0.30 per cent. Both 85-5-5-5 and 88-10-2 precipitate delta. The specification limit of usage is 406°F., temperature of saturated steam at

Jan 1, 1945

By J. Rogers, K. L. Sutherland

THE relationships between collector and mineral, activator and mineral, and activator, collector and mineral will be considered herein. We propose to criticize current theories of flotation but we will not repeat the many pertinent arguments presented by Wark and Cox28 in 1937 in which the "chemical" or more strictly the "solubility" theory proposed by Taggart and co-workers22 is shown to be untenable. Recently Taggart, Arbiter and Kellogg24 have studied the long-chain aliphatic amines as collectors and have attempted to resuscitate the solubility theory for these particular collectors. Dean and Ambrose3 have felt that "we stumble over words" when we distinguish different mechanisms of adsorption. Our view is that nothing is more stultifying to the development of a science than attempts to force widely different phenomena into one classification. This oversimplification leads immediately to difficulties in the interpretation of results. EXPERIMENTAL PROCEDURE The response of minerals to activator and collector was determined by "captive-bubble" tests. The methods of preparation of specimens26 and of testing18 have been previously described. Results from captive-bubble tests were confirmed by flotation tests in glass-stoppered cylinders.25 The full lines shown in the figures relating solution composition to contact (flotation) distinguish solutions in which contact (flotation) is possible from those in which it is not. The curve is derived by successive approximation and the accuracy is ±0.2 pH units and ± 5 mg. per liter of collector or depressant. Solution compositions for which it was shown experimentally that contact or flotation is possible are marked by a circle, those for which it is not possible by a cross and those for which it is doubtful by a cross within a circle. It is assumed that at all other solution compositions within (without) the curve, contact or flotation is possible (not possible). After testing in the collector solution, the mineral was removed, washed, and tested by the captive-bubble method in water, or floated in a cylinder with frother only. In this way the composition of solutions in which a mineral was conditioned but in which properties of the collector solution prevented contact or flotation can be defined. The composition of solutions in which conditioning was possible lie between the full and broken curve, the experimental results being indicated on the figures by a square. The curves show the limits of composition of solution in which flotation in a cell would be possible. Many factors other than those considered in these tests determine recovery or grade of a given mineral in practice.

Jan 1, 1947

By Jack L. Taylor

WITH the exception of the work of Vogel,&apos; and Rolla and Iandelli, very little information has appeared in the literature on titanium binary systems with the rare earth elements. Rare-earth additions to titanium alloys may possibly parallel the beneficial effects on the physical properties of certain steels. Having this in mind, some knowledge of solubility limits and intermediate phases, if any, would be useful. An investigation of the titanium-rich Ti-Ce system was started. Alloy Preparation The iodide titanium from the New Jersey Zinc Co. and the cerium metal from the Institute for Atomic Research Ames Laboratory used in this work were supplied with the typical analysis shown —iodide titanium: 0.0065 pct Mn, 0.0022 pct Fe, 0.0015 pct Cu, and 0.0042 pct Pb; cerium metal: <0.01 pct Fe, <0.03 pct Si, <0.05 pct Ta, <0.02 pct Pr, <0.03 pct Ca, <0.02 pct Mg, with no trace of Nd, La, or Th. The cerium metal, stored in oil, was washed and scraped just before weighing into capped cups machined from iodide titanium. The cold hearth helium-arc furnace described by Schramm et al3 was used for melting. The alloy buttons from the arc-melted cup were cut in two, mounted, and examined. Homogeneous alloys were demounted and rolled to approximately 1/4-in.sq bars. Sections l/8 in. thick were cut from the bar to provide samples. The Ti-Ce alloy buttons analyzed for cerium were of the following composition: 0.19, 0.36, 0.48, 0.61, 1.15, 2.52, 3.06, 3.70, 4.30, 9.20, 14.26, and 18.50 wt pct. Nominal

Jan 1, 1958