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By R. R. Bryan

SINCE the first applications of cyanidation to silver ores about 1906, treatment of ores in the Pachuca district has been entirely by straight cyanidation. Until about the year 1921, Real del Monte removed a sulphide concentrate (largely pyrite) by tabling in the cyanidation grinding circuit. The ores milled prior to the year 1935 were dry ores, ideal for cyanidation, and yielded 90 pct of the silver and 92 pct of the gold. Since 1934 there has been a notable increase in the refractivity to cyanidation of the average ores received by the mill. Tonnages of the former amenable ores have gradually been replaced by tonnages of more refractory ones, so that present average recoveries have fallen to 83 pct of the silver and 90 pct of the gold. During this period there has also been an increase in the content of base metal sulphides' in many of the ores although this increase is not always directly related to the increased refractivity. Since March 1950, the Cia. de Real del Monte has been developing a flotation treatment for some-of its ores before cyanidation. Suitability for flotation is judged not only by the base metal content, but also by the amount of additional recovery of silver and gold made possible by the combined treatment. Selection and segregation of ores most suitable for flotation before cyanidation is a difficult problem for the mines, and the best that can be done is a compromise that leaves much to be desired. Development of Pre-cyanidation Flotation at Real del Monte Much laboratory work has been done on flotation of ores from the Pachuca district, however, the data could not supply reliable information on the many factors connected with the economics of low-base metal-high silver ores. Pilot plant operations would be required. Pilot plant facilities were made available in March 1950, when it became necessary to close down Loreto's 100 ton per day selective flotation plant because of exhausted ore supply. The results of the 100-ton pilot plant tests indicated that a large part of our tonnage of cyanidation ores could be profitably treated by flotation, and a tonnage of 1100 tons per day was selected as the maximum tonnage of ore that the mines were likely to deliver, and because this tonnage fitted with the grinding capacity of this circuit. Equipment was ordered for the production of a single flotation concentrate. The conception of flotation before cyanidation, at that time, was the production of a single concentrate to be shipped to the lead smelter. The 100-ton pilot plant was immediately increased to 350 tons by installing additional flotation cells.. This. plant was operated continuously until the 1100ton plant was put in operation in January 1951. In July 1950, an increase in the quotation for zinc called for an investigation- of the possibilities of producing a zinc concentrate, and laboratory tests on the tailings from the 350-ton plant indicated that a

Jan 1, 1952

By F. O. Garrabrant

POWER is furnished to the Climax Molybdenum Co. by the Public Service Co. of Colorado over two 100,000-volt line to a bank of three 3333-kva. Transformers 100/13.8 kv. These transformers arc so designed that with the addition of forced cooling their capacity may be increased 33 1/3 per cent at 12,000-ft. altitude. Under present conditions, these transformers are operated at 140 per cent of their rating with forced air cooling. On the secondary side of this bank there is a 15-kv., 1200- amp. Westinghouse oil circuit breaker provided with automatic reclosing relays allowing three successive short circuits before lockout. From this breaker 13.8-kv. power is transmitted to a central switching station (Fig. I) where all circuits radiate to the different operations of the property. Each circuit is protected by a 23-kv., 600-amp. oil circuit breaker with by-pass and isolating disconnecting switches. MINE Power at 13.8 kv. enters the mine through a ventilation raise over a three-conductor, size 1-0 varnished cambric insulated, leaded, steel-wire armored cable, insulated from 17 kv. This power is received and distributed through one switching station and four underground substations, shown in Fig. 2. No. 370 substation (Fig. 3) contains three 250-kva, 13,800/440-volt, single-phase Pyranol transformers protected on the high-voltage side by a 600-amp. manually operated oil circuit breaker. This station serves 20 slusher-hoist motors, a 45-kw., three-phase, 220-volt electric furnace, a Massco hot miller for tempering and milling Timken detachable bits and a hoist with 175-hp. motor on the No. 3 development shaft. Also at this shaft are three pumps. One two-stage pump, 1200 gal, per min., 100-ft. head, 3600 r.p.m., driven by a 150-hp. motor, is at the bottom of the shaft. Two deep-well pumps, 500 gal. per min., 500-ft. head, are at the collar of the shaft. Under normal operation the one pumping 1200 gal. per min. is controlled by a float switch to take care of the water. If this pump fails, or if its capacity is insufficient, the first 500-gal. pump is automatically started, and if still more capacity is needed the second 500-gal. pump automatically starts. Time delay relays are provided on the deep-well pumps, so that if power fails the column will have time to drain before the pumps restart. On failure of any of these, an alarm is sounded in the train dispatcher's office, so that mechanics can be sent to ascertain the cause of failure and make repairs. The cable from No. 370 substation is run to the switching station, in which are four15-kv., 600-amp. direct-current solenoid-operated oil circuit breakers. The current for operating and tripping is furnished by a 125-volt storage battery. All breakers are operated from the switchboard in the compressor room. One circuit from this switching station goes to the

Jan 1, 1944

By L. E. Young

LOADING machines may be classified in several ways: (1) Machines which cut or break down and load .the coal; (2) machines which simply load the coal; (3) devices which load and transport the coal; (4) de-vices which elevate; (5) devices which transport the coal after it has been shoveled onto the device. There are many types of mechanical loading devices, Bull. 17, Coal Mining Investigations, under the auspices of Carnegie Institute' of Technology, U. S. Bureau of Mines, and the Advisory Board, listing 31 exclusive. of -the large shovels used in thick coal seams. Mechanical loading of coal is not in its infancy; I think some of the early patents go back to 1890. One large coal company, properly called the early leader in this field of mechani-zation, has about 26 large machines at work. This company spent about thirty years in developing a ma-chine, and at the present time is loading approximately two million tons of coal a year in its mines. The chair-man directed a question at the previous speaker; inquir-ing if real progress was being made in mechanical load-ing. The experience of the company just referred to is in itself an adequate answer. A review of a list of mines using mechanical loading devices shows that there are 28 mines which we may say are 100 per cent mechanical. There are four more mines which have been 100 per cent mechanical, but which are idle at this time on account of labor condi-tions. There are at least 25 more at which the tonnage is more than 50 per cent mechanical. In speaking of a 100 per cent mine I do not mean that every pit car of coal is loaded by machine, because when complete, re-covery of the coal is attempted it is not practical to load all the coal at, the break line. In order to load the last 5 to 10 per cent of coal with a minimum of gob material and in order to keep the. working places along the break line timbered in a safe manner it is generally desirable to use a few hand loaders. At a certain mine operat-ing on the block system 90 percent of the coal is being loaded by what is properly a mechanical loading: device, and about 10 per cent -is being loaded by hand. This 10 percent is right off the break line, out of what might be called, the screen pillar. The mines referred to are scattered over the country, . in seven different' States, and the equipment used is produced by ten different machinery companies. As to size, these mines vary from small mines of 500 tons mining capacity-up. In one instance a company with one loading machine is producing an, average of 450 tons per day. One of the large mines that is idle pro-duced at the. rate of 4300 tons per day, of which 97 per cent was actually loaded by machinery.

Jan 8, 1928

By R. F. Ashton, D. S. Thompson

In reporting results of precipitation hardening experiments, it is customary to include such conditions as solution heat-treatment temperature, specimen size, and quench medium as well as the aging time and temperature. One portion of thermal history that is often overlooked is the effect of the heating rate employed to achieve the aging temperature. It is the intent of this note to demonstrate the importance of the rate of heating to aging temperatures. This work was carried out on the aluminum alloy 7075 aged at 162oC, a temperature commonly used to obtain stress corrosion resistant tempers. A 2 in. thick plate of 7075 was quenched into 24°C water after solution heat-treatment for 2 hr at 470°C. Approximately 13 in. around the periphery of the plate was discarded to minimize quench rate variations, and the remainder was divided into 0.5 by 1.8 by 2.0 in. blocks which were held at room temperature, 24oC, for 4 days. This aging, or incubation period, at room temperature allowed a copious supply of GP-zones to form. Blocks were then heated at various rates to 162oC using the practices shown in Table I. Both single- and two-step aging practices were employed for each heating rate. The single step consisted of heating directly to temperature at the indicated rate while the two-stage practice included an interruption of heating for 7 hr at 107°C. In the case of the two-stage 100°C per min heating rate experiment, the specimens were returned to room temperature after 7 hr at 107°C where they were held for 16 hr be-fore finally heating at this rate to 162°C. This devia-tion in practice is not considered to be significant. Blocks were removed after various times at 162oC, and aging curves were determined from duplicate R. F. ASHTON and D. S. THOMPSON,Member AIME, are Senior Scientists, Metallurgical Research Division, Reynolds Metals Company, Richmond. Va. Manuscript submitted December 2. 1968. IMD Table I. Aging Practices Two-step Single-Step Heating Rate* Aging Aging Heating Method oC per Min A B Programmed Air Oven 0.23 * 0.01 C D Programmed Air Oven 0.93 ? 0.03 E F Air Oven at Temp -10 G H Oil Bath at Temp ˜100 *Heating rate was measured with a thermocouple at the center of a specimen. An average value between the starting temperature and 3°C below the aging temperature is reported for practices E-H. PRACTICES 70 — A B C E and G 50 - 2 4 6 810 20 40 60 HRS.AT 162°C Fig. 1—The influence of processing variables on 7075 aging curves. Fig. 2—7075 aged according to practice H (single-step aged, fast heating rate) for 3 hr. Magnification 100,000 times.

Jan 1, 1970

By L. D. Jaffe

IN previous metallographic work on titanium and its alloys, difficulty has been encountered in distinguishing spheroidal particles of titanium hydride, dispersed in a-titanium, from other phases that may be present, such as p-titanium. This problem is common in examination of commercial unalloyed titanium.' In the course of an investigation of the Ti-H phase diagram, it was noted that titanium hydride, dispersed in a matrix of a-titanium, showed strong optical anisotropy when examined under polarized light. With sensitive tint illumination, particles of the hydride, when the stage is rotated, changed from bright yellow to bright blue in color. Since P-titanium is cubic, it showed no color change on rotation, remaining dark gray or blue-gray. Although a-titanium is hexagonal, it showed only a minor color change, going from pale pinkish-blue to pale purplish-blue. These colors could be modified somewhat by adjustment of the illumination. They were observed in specimens mechanically polished, both without etching and after etching with hydrofluoric acid-nitric acid mixtures in either water or glycerine. The strong color change of the hydride seems to offer a simple method for its metallographic identification. Optical anisotropy of titanium hydride is inconsistent with its accepted face-centered-cubic structure.' Samples of iodide titanium containing 6 to 40 atomic pct deuterium were prepared for diffraction studies. After deuterating and slow cooling' the samples were held 21 % hr at 400 °C, water quenched, held 61 hr at 100°C and quenched, then held 64% hr at 255 C and quenched. Debye ex-posures with MoKa X-rays and 1.111A neutrons were made at room temperature of samples as-deuterated and after the 100" and 255°C treatments. Table I shows the lines observed, in addition to those attributable to a-titanium. These data show that the hydride phase was not cubic, but probably tetragonal. The titanium atoms appeared to form a body-centered-tetragonal lattice complex, two titanium atoms per cell, with approximate dimensions: a = 3.12A, c = 4.18A, and c/a =1.34A. This is equivalent to a face-centered-tetragonal complex, four titanium atoms per cell, with approximate dimensions: a = 4.42A, c = 4.18A, and c/a - 0.946A. The presence of the 2.98 and 1.81A neutron lines indicates that the true Bravais lattice is not identical with this lattice complex, but may be primitive tetragonal with dimensions some multiple of those of the complex. The relative intensities of the hydride and a-titanium X-ray lines as functions of overall composition indicated that the hydride, presumably in equilibrium with the a at 100" to 255"C, contained more than 62 atomic pct H. Acknowledgments The experimental work covered in this report was done at Brookhaven National Laboratory, through the courtesy of D. Gurinsky. The cooperation of J. Sadofsky, 0. Kammerer, A. McReynolds, and A. Cendrowski is gratefully acknowledged.

Jan 1, 1957

By Gordon Fletcher

Pacific Rim countries currently have plans to import about 100 Mt/a (110 million tpy) of steam coal by 1990. Due to its close proximity, Powder River Basin coals offer a long term economic and dependable source of supply to the Pacific Rim. In existing boilers, Powder River Basin coals can be blended with Australian, South African, or Canadian coals. For new units, Powder River Basin coals are very competitive on a total evaluated cost basis. The advantages of using Powder River Basin coal include: assurity of supply, consistent coal quality, stable labor force, and attractive mining economics. (NOTE: PAPER NOT RECEIVED IN TIME FOR PUBLICATION.)

Jan 1, 1982

Magazine, Advertising $ 32,823 60 Sale of Magazines 3,93329 Total Magazine $ 36,756 89 Dues, 'Arrears 5,866 97 Dues Current 100,223 80 Dues of New Members 6,143 29 Dues in advance 1,586 32 Initiation fees 6,820 00 Sale of Transactions 5,464 99 Sale of Binding 5,618 90 Sale of Pyrometry Volume 336 30 Sale of Indexes and Emmons 106 10 Sale of Pamphlets 575 32 Miscellaneous Receipts 105 60 Interest on Bank Balances 1,217 67 Interest on Life Membership Investment 512 18 Interest on James Douglas Library Fund 5,678 67 Total Receipts $177,013 00

Jan 1, 1923

By C. P. Kempter

AS part of a continuing program of liquid metals research, the densities of liquid lead, tin, zinc, cadmium, and indium have been measured by a pycnometric method. The pycnometers are made of high-density graphite to the design illustrated in Fig. l(a). In pursuing this research program, emphasis has been placed on detailed experimental care. As a result, claims are made for densities accurate to within ± 0.05 pct over temperature ranges varying from about 100± to 200°C above the melting point. A paper reporting the densities of tin, lead, and Pn-Sn alloys has recently been published,' while papers dealing with

Jan 1, 1969

By A. F. Crowley, D. W. G. White

AS part of a continuing program of liquid metals research, the densities of liquid lead, tin, zinc, cadmium, and indium have been measured by a pycnometric method. The pycnometers are made of high-density graphite to the design illustrated in Fig. l(a). In pursuing this research program, emphasis has been placed on detailed experimental care. As a result, claims are made for densities accurate to within ± 0.05 pct over temperature ranges varying from about 100° to 200°C above the melting point. A paper reporting the densities of tin, lead, and Pn-Sn alloys has recently been published,' while papers dealing with

Jan 1, 1969

(Under this heading will be published notes sent to the Secretary of the Institute by members or other persons. ) Position of laboratory assistant in metallurgy, pyrometry, and physical testing is vacant. Salary $80 to $100 to start. Knowledge of metallography desirable. Opening for a mine superintendent familiar with stoking methods on narrow gold quartz veins; also for a. superintendent of a 50-ton fine-grind-in, cyanide mill. College man under 40 preferred. Position in sales department is open for a mining engineer, with experience in mine ventilation and possessing some knowledge of fan design; one competent to lay out ventilating systems to meet various conditions.

Jan 10, 1913

By Fortunato Pereira Gamba

The Department of Narino (formerly included in the Department of Cauca) is a narrow band about 170 km. (100 miles) wide in the southern part of the Republic of Colombia, near the boundary of Ecuador, as showu in the sketch-map, Fig. 1. Its capital is Pasto, a small city of 25,000 inhabitants; the pri~icipal towns are T?querres, Cumbal, Ypiales, Barbacoas, and Tumaco. The port on the Pacific ocean is Tumaco, and traffic in the interior is carried partly on the Patia river (steam navigation to the fluvial port of Barbacoas), and afterwards overland over poor roads by human bearers or on mule-back.

Jan 1, 1913

By R. W. Smith, K. H. Lee

Bond type simulated closed-circuit grindability tests and batch grindability tests were run on six pure minerals plus a limestone and a taconite ore. The simulated closed-circuit tests were made at five different meshes of grind. The minerals tested varied widely in hardness and breakage pattern. For all simulated closed-circuit tests both Bond and Holmes work indices were calculated and compared. The Bond grindability testlm6 has been widely and quite successfully used for predictions of such items as ball mill power requirements and as an aid in plant comminution equipment selection. The test is interesting in that although run as a series of batch tests it is over-all a simulated closed-circuit test with a circulating load of 250%. The simulated closed-circuit feature is probably the primary reason for the notable success of the test. Perhaps the major criticism that can be brought against the test is that it is time consuming, since it normally takes a trained technician about an eight-hour shift to perform it.4 For many purposes this is an acceptable time requirement. However, if a grindability test is to be used for mill control purposes it should be performed in as short a time interval as possible, preferably under an hour. Sometimes in industry strictly batch tumbling mill tests are run and an attempt is made to correlate such batch tests with plant mill operation. Then an attempt is made to adjust operating tumbling mills based on data from the tests. The tests, however, frequently do not correlate well with operating data. The purpose of the present work was to explore some of the similarities and differences between a simulated closed-circui.t grindability test and a batch type test. Six different minerals plus a limestone and a taconite ore were tested in the experimentation. These materials varied widely in hardness, breakage characteristics, etc. Standard dry Bond grindability tests were mn on all materials at 48, 65, 100, 150 and 200 meshes of grind. Feed material was prepared by stage cmshing to -6 mesh. Initial feed volume was 700 cc. Care was taken to maintain the standard constant weight (about 20125 gm) and number of balls (283) during all the tests. Rotational speed was 70 rpm. Table I lists the materials studied along with pertinent data regarding the materials and Bond grindabilities obtained. In addition, dry batch tests were performed on all materials. These tests were made in the same mill as was used for the Bond tests. An identical ball charge and an identical feed volume were used. After rotating the mill for a predetermined number of revolutions at the same rotational speed as for the Bond test, the material was removed and separated from the ball charge. A screen analysis was performed on this product. Knowing the original size distribution of the feed and the size distribution of the product, it was possible to calculate the grindability or grams of a particular mesh produced per revolution of the mill. After the screen analysis of the product, the various fractions were recombined and placed back in the mill for further grinding. After grinding for additional revolutions of the mill the procedure was repeated. Thus, it was possible to obtain grindabilities for each material at a number of meshes of grind and for various revolutions of the mill. Grind-abilities obtained from batch tests are shown in Table 11. COMPARISON OF BOND AND BATCH GRINDABILITIES Figs. 1A through 1D show relationships between batch and Bond grindabilities. Figs. lA, lB, 1C and 1D show respectively graphs indicating the relationships for batch tests run for 100, 300, 500 and 700 revolutions of the mill. For a particular material on each graph, points are shown for 200, 150, 100, 65, and 48 meshes of grind except for several missing coarse meshes of grind for galena which did not fall conveniently on the figures. For each material and graph the point nearest the origin is for 200 mesh, the next for 150 mesh, the next for 100 mesh, etc. Considering Fig. lA, it is interesting to note that all relationships shown are essentially straight lines. There are, however, considerable differences

Jan 1, 1969

By R. J. P. Lyon, A. E. Prelat, H. Lawrence

Introduction In 1977, the US Bureau of Mines awarded a grant to the state of South Carolina to explore the possibility of using Landsat multispectral scanner (MSS) data as an aid in monitoring surface mining activities in selected counties of the state. Work had been done in Aiken County prior to the grant by USBM and state personnel at the EROS Data Center using both the GE Image 100 and ESL IDIMS systems. The initial work on the 1:100 consisted of locating and measuring the areas of various kaolin mines. The results were excellent on location, identification, and accurate area measurements of several mines as compared with aerial photos of each mine. The effort also included locating and measuring the size of all mines (including sand and gravel) within two US Geological Survey (USGS) 7 % - min quadrangle maps. Line printer overlays using the IDIMS equipment and software resulted in the location of all known mines within an estimated error less than zt 91.4 m. Mine acreage was calculated by counting each symbol (which represented one pixel) and multiplying by 1.1 (the number of acres per pixel). [Fig. 1] shows a portion of one quadrangle overlay. [Table 1] gives a comparison of acreage measured for several mines from the 1-100 and IDIMS system vs. aerial photo measurements. The success of the preliminary work resulted in the South Carolina Land Resources Conservation Commission submitting a proposal to USBM for funding to expand the scope of work to other mineral mining operations within the state and determine if Landsat Imagery could be used to accurately locate and measure mining and reclamation activity on a continuing basis. Three separate tasks were proposed for selected kaolin mines in Aiken County: Task 1, to measure changes in active mine developments over the period 1974-77; Task 2, to measure reclamation activity, and, if time permitted, seasonal changes in mine signatures over the same period of time; and Task 3, to classify one complete mining operation (i.e., differentiate between active, spoil, overburden, water, reclaimed portions and surrounding terrain. The major effort was centered around J.M. Hubeis Richardson mine. Mining operations were progressing in a northerly direction and included three areas of reclamation: one prior to 1974; one started during early 1974; and one reclaimed during 1976. A discussion of all work done under the grant is available in the March 1979 Final Report listed under References. Computer compatible tapes (CCTs) containing MSS data from Landsat Path 18, Row 37, were purchased from the National Aeronautics and Space Administration (NASA) covering the period February 1974 through January 1977. The MSS data were processed on the General Electric Image 100 using parallelepiped classification at the EROS Data Center's Data Analysis Laboratory, Sioux Falls, SD, and at the Stanford Remote Sensing Laboratory, Stanford University, Palo Alto, CA, utilizing both an unsupervised classification (ISOMIX) and a supervised classification (sequential discriminant analysis). Studies at Stanford University Remote Sensing Lab (SRSL). In Feburary 1977, initial work was begun at Stanford using the STANSORT computer program running on a DEC PDP-10 (Honey, Prelat, and Lyon, 1974). This software is fully interactive and designed to study Landsat CCT data at high resolution (small areas) on a pixel-by-pixel basis (Lyon, 1977). Unsupervised Classification. The initial step was to locate the Richardson mine on the positive print of the full Landsat scene from CCT 1644-15252, dated 6/17/74. This was relatively easy due to a large abandoned kaolin mine and a long narrow pond clearly visible in the scene. Two unsupervised classifications of the Richardson mine were run using ISOMIX clustering (Honey, Prelat and Lyon, 1974). (ISOMIX essentially follows the interactive clustering procedures of ISOCLS Kan, Holley, and Parker. 1973). Although the classifications were successfully accomplished, it was apparent from the TV display that the results did not clearly define the mine areas. A supervised classification would have to be made in order to clearly separate each portion of the mine. Supervised Classification. This classification was done using the Stepwise Discriminant Program (BMDO7M) which is part of the BIOMED package (Dixon, 1970) and involved the following steps: Training group populations were selected and the digital

Jan 1, 1982

By R. P. Aikman, F. G. Miller, J. M. Podgursky

As part of a larger study of the mechanism of froth flotation of coal, the rates at which coal particles of various size and specific gravity react to froth flotation were investigated. Results of this study show that in flotation: I) recovery rate increases, but selectivity decreases with a decrease in particle size down to 100 mesh, and 2) recovery rate is constant for all sizes between 100M and 325M. n response to the growing demand for low-sulfur coking coal for iron making, the coal industry is continually striving to perfect methods for cleaning coal to acceptable ash and sulfur levels while maintaining maximum product yield. By present standards the industry has adequate processes for cleaning coarse coal, i.e., + 28 mesh, at almost any desired specific gravity. However, in the case of fine coal, i.e., —28 mesh, neither of the two existing processes - froth flotation or hydrocycloning — in its present state of development yields products that are comparable in ash and/or sulfur content to products normally obtained from coarse coal processes. The work outlined in this paper is part of a larger study directed toward maximum exploitation of froth flotation for fine coal processing. The concepts and data presented are now serving as a basis for designing fine-coal circuits in which improved efficiency is obtained by optimum use of froth flotation with other fine-coal processing methods. The paper includes: 1) A discussion of the value of a detailed wash-ability study as a method of assessing the problems associated with fine coal processing, 2) A general discussion of hydrocyclone principles, included to serve as a basis for understanding the mechanism of coal flotation, and 3) A detailed discussion of the mechanism of coal flotation, as well as the research leading to the optimum use of froth flotation in combination with other fine-coal cleaning methods to provide efficient fine-coal processing. Our study of the rates at which coal particles of various size and specific gravity react to flotation shows that in flotation: 1) recovery rate increases but selectivity decreases with a decrease in particle size down to 100 mesh, 2) recovery rate is constant for all sizes between 100 mesh and 325 mesh, and 3) the flotation process effectively rejects only those extremely fine particles which do not contain surface carbon. (Because of this feature, froth flotation is an excellent method for rejecting clays - the only extreme fines in - 325-mesh coal that are always carbon free.) Washability studies of flotation products are used to assess flotation efficiency. These studies show that: 1) the increase in percent sulfur in flotation products above the theoretical sulfur content as defined by the washability study is due almost entirely to flotation of fine high-specific-gravity particles that are part coal and part pyrite, and 2) in a given size range particle floatability decreases with an increase in particle specific gravity. Error curves are explained and are used to compare the mechanisms of coal flotation and hydrocycloning. These comparisons show that hydrocyclone performance is affected primarily by particle mass, i.e., particle specific gravity and size, whereas coal flotation performance is dependent on both the particle mass and the coal surface available for bubble attachment. Since the error curve is a basic tool for rating the efficiency of any coal processing system, it will be useful to quickly review the principles of error curve construction. RATING COAL-PROCESSING EFFICIENCY BY THE ERROR CURVE The organic component of coal is of lower specific gravity than its impurities, and the most effective principle for coal concentration is therefore separation on the basis of specific gravity. However, since there is no sharp dividing line between pure coal and pure ash components, specific gravity must be closely controlled if the product is to have acceptable quality at maximum yields. In turn, the basis for close control in any coal processing system is the reference standard provided by precise laboratory washability studies. In these studies, coal is carefully separated into a series of progressively higher specific gravity fractions to establish a relationship

Jan 1, 1968

By James L. Poole

INTRODUCTION Companies in the mining industry are subjected at times to currency exchange rate risks and interest rate risks. The former occurs any time a firm deals in more than one currency. The latter occurs anytime a firm is financing itself with borrowings that have a floating interest rate (although a fixed rate can be a problem at times). FOREIGN EXCHANGE RISKS There are two general reasons why exchange rate risks (FX risk) can occur: First, a firm may be building a mine or plant in another country where financing for the plan is denominated in funds other than the country hosting the plant. For example, if a Canadian firm is building a smelter in another country which is being financed with Canadian dollars, the C$100 million for labor, which must be paid in the local currency, has subjected the firm to a C$100 million FX exposure risk. If after the cost estimates were made the local currency appreciates 30% against the Canadian dollar, the cost of labor has increased to C$130 million. (Of course, the local currency could also devalue during this period making the labor cost component less than estimated.) The other reason exchange rate risk arises is due to the fact that the costs of production in one country producing minerals may be incurred in one currency while the production is sold in another currency, such as coal being exported from Canada to Japan under contracts denominated in Yen. In this example, the Canadian firm is exporting coal to Japan and being paid in yen in 30 days at a contract price of 100 Y per ton. Furthermore, assume that the mining firm anticipates a depreciation of the yen against the Canadian dollar of 5% in the next 30 days. Therefore, the mining firm stands to lose 5% of the value of its shipment in a month's time not including carrying costs. In both cases, the mining company risks an adverse change in the exchange rate, that is, the currency the company holds may decrease in value relative to the other currency. There are a number of ways to mitigate the exposure to foreign exchange which can be utilized by the mining company. Optimally, a company should try to match all their inflows and all their outflows in the same currency. For example, if a Canadian firm has mining costs and debt service costs both denominated in Canadian dollars then 100% of their receipts would need to be in Canadian dollars. If a Canadian firm has 70% of its cash outflows in Canadian dollars and 30% of its cash outflow in American dollars in the form of a loan amortization then their receipts would need to be 70% in Canadian dollars and 30% in U.S. dollars. This general principle of matching cash receipts and cash costs in the same currency works fine in a perfect world but what of the real world when these cannot be matched? In actual practice, however, the financial markets have developed a number of ways in which foreign exchange risks can he managed or reduced. FORWARD MARKETS One method is to use the forward markets which are made on a world-wide basis by the commerical banks for foreign exchange. If a firm wanted to lock in the current exchange rate for the receipt of a foreign currency at some point in the future, they could do so by contracting on the forward market with a bank selling that foreign currency at the current rate at the future date. For example, assume that a Canadian firm is exporting coal to Japan which would pay yen in 90 days. If a firm wanted to lock in today's exchange rate they could contract with a bank to sell on the forward market yen in 90 days. When the yen was received the contract would be executed by selling the yen to the bank and receiving the previously agreed upon number of Canadian dollars. Generally speaking, the forward market can be used to sell forward about 12 months. The costs of selling (or buying) forward

Jan 1, 1985

All the TECHNICAL PUBLICATIONS and CONTRIBUTIONS published in 1940 are available at Institute headquarters, unless otherwise noted. They are also on file in many public, university and technical libraries and, when so indicated in the list, may be found in the TRANSACTIONS. TECHNICAL PUBLICATIONS in Classes A-Metal Mining, B-Milling and Concen-tration, H-Industrial Minerals, and I-Mining Geology are distributed in MINING TECHNOLOGY, which is issued every other month. Technical Publications in Classes C-Iron and Steel Division, D-Nonferrous Metallurgy and E-Institute of Metals Division are distributed in METALS TECHNOLOGY, which is issued twice quarterly. TECHNICAL PUBLICATIONS in Class G-Petroleum Division are distributed in PETRO-LEUM TECHNOLOGY, which was issued quarterly in 1940 but will be issued every other month in 1941. TECHNICAL PUBLICATIONS in Classes F-Coal Division, J-Mineral Industry Education Division, K-Mining Administration, and L-Geophysical Methods of Exploration are issued as separate pamphlets. An Institute member is entitled to one copy of any TECHNICAL PUBLICATION or CONTRIBUTION. Copies of TECHNICAL PUBLICATIONS listed have been mailed to members of the Institute accord-ing to their registration by classes. CONTRIBUTIONS and other TECHNICAL PUBLICA-TIONS will be sent upon request. A member may purchase an additional copy of a TECHNICAL PUBLICATION or a CONTRIBUTION at one cent a page; minimum charge, 25¢. Nonmembers may purchase any CONTRIBUTION or a TECHNICAL PUBLICATION of the F, J, K, or L classes at the same rate, but the nonmember price per copy for METALS TECHNOLOGY, MINING TECHNOLOGY or PETROLEUM TECHNOLOGY is $1.25 domestic, $1.35 foreign countries. TECHNICAL PUBLICATIONS (separate pamphlet form), $2 per annum domestic, $2.50 foreign. METALS TECHNOLOGY, $7 per annum domestic, $7.80 foreign (Iron and Steel Division and Institute of Metals Division ferrous and nonferrous papers). MINING TECHNOLOGY, $6 per annum domestic, $6.60 foreign (metal mining, milling and concentration, Industrial Minerals Divi-sion, and mining geology papers.) PETROLEUM TECHNOLOGY, $6 per annum domestic, $6.60 foreign. Bound volumes have been mailed to members that have requested them, and MINING and METALLURGY is mailed each month. A member is entitled to one volume a year on his membership and he may purchase for $2.50 one copy of a volume not already received; additional copies of the same volume, $5. Nonmembers may pur-chase the volumes listed, at $5 each, plus 60 cents for Canadian or foreign postage. Extra single copies of MINING AND METALLURGY, 25¢ current issue, back numbers 50¢ each to members; to nonmembers 506 current issue and $1 each, back issues. Subscription price to nonmembers and to members for extra copies, $3 per annum domestic, $4.00 foreign.

Jan 1, 1940

By H. H. Klepfer, M. E. Snyder

UNTIL very recently, the commercial availability of the rare earths as metals has been very limited. Fabrication of mill products from these metals has not been studied in most cases. This note reports the results of the development of fabrication techniques for thulium. Thulium has a melting point of about 1550°C and a hexagonal close-packed crystal structure. It oxidizes in air to give a black oxide (TmzO,). The procedures for producing thin thulium foil were developed on one ingot weighing about 220 g and were subsequently applied in processing about a pound of metal to foil. Several alternates for the various fabrication steps were investigated and will be discussed. The metal fabricated was in the form of commercial chill-cast ingots 1 in. in diam and 2 in. long weighing approximately 220 g. Impurities in the ingots were reported by the vendor to be 4000 ppni tantalum, 2000 ppm calcium, 200 ppm nickel, 100 to 200 ppm iron, 100 ppm europium, and less than 100 ppm copper, lutetium, and ytterbium. In addition to these impurities, several salt-like inclusions as large as l/8 in. in diam were revealed along the center line of the one ingot sectioned. Preliminary tests indicated that small wafers cut from the as-cast ingot would not fabricate readily by rolling. Forging of copper-jacketed wafers was therefore attempted. At 1550"F forging was satisfactory but an apparent reaction of copper with thulium demanded investigation of lower temperatures. Therefore, the remainder of the test ingot was forged at 1450°F—with only minor cracking. All ingots forged were inserted into copper tubes of 1 in. 1D and 0.125-in. wall thickness. The jackets were sealed by flattening the ends of the tubes and welding under helium. Heating time at 1450°F was 30 min. Press forgings of 1/8 in. per pass were used, followed by 10 min reheats. When the ingots had been squared and reduced to 0.250 in. in thickness, the original copper jacket was stripped off and replaced by a new jacket in preparation for hot rolling. Hot rolling at 1450" F without edge cracking was readily accomplished after forging. Excellent results were obtained with 10 pct reductions of thickness followed by 5 to 10 min reheats. After reduction of thickness from 0.250 to 0.100 in. the copper jacket was removed. It was found, in fact, that hot rolling in air was possible. A tenacious black oxide similar to that seen on zirconium was formed during 3 min reheats at 1450°F. Reduction in air to 0.010 in. foil was possible taking 10 pct reductions per pass. The oxide coat formed during hot rolling in air could best be removed by sand blasting and pickling. Common pickling solutions containing polar solvents were found to attack the metal too rapidly and a concentrated nitric-hydrofluoric acid mixture attacked neither the oxide nor the metal. The most satisfactory pickling solution was 52 vol pct concentrated nitric acid-48 vol pct glacial acetic acid. After forging to 0.250 in., vacuum annealing and cold rolling was found to be another satisfactory alternate to hot rolling in copper jackets. After forging. a hardness of Rockwell B76 was found. Annealing in vacuum (2 X l0-5 mm of Hg) for 1 hr at 1200"F did not alter this value. Annealing at 1470"F for 1 hr brought the hardness down to R;]63. With cold rolling (5 pct per pass) the hardness returned to about R,1:76 after 20 pct reduction and edge cracking became noticeable. However, cold rolling to a total of 40 pct reduction in thickness (RI,,83) was possible before edge crack propagation became serious. Good surface finish was obtained, and the metal loss due to oxidation was minimized by cold rolling and vacuum annealing. Using this procedure the yield of 1.25 in. wide by 0.010 in. thick foil from a 1-in. diam ingot was about 40 pct.

Jan 1, 1961

By A. G. Guy

Thermodynamic activity rather than chemical composition is basic to the analysis of diffusion. This is the essential conclusion reached by Darken1-3 and by Birchenall and Mehl.4 If so, it is reasonable to expect that diffusion calculations should be carried out using activities. This paper will illustrate a method of employing activities in analyzing interstitial diffusion. For interstitial diffusion (of carbon in iron and probably generally) the diffusion of the solvent atoms, volume changes, and other such complexities not explicitly provided for in the usual diffusion equations cause a negligible error of the order of 1 pct. Therefore it is relatively easy to test the validity of an analysis based on activities in the instance of interstitial diffusion. The mathematical treatment of the more complex substitutional diffusion could be simplified once the adequacy of the activity method had been established for substitutional diffusion. Thus the present work has the purpose not only of providing a superior analysis of interstitial diffusion but also of indicating a possible approach to the more important question of an adequate treatment of substitutional diffusion. Development of the Diffusion Equation Although activities are convenient quantities to determine experimentally, they are derived from the more fundamental fugacities.9 Possible confusion in using activities can be avoided by basing equations involving their use on the relationship, dP = -D1d/ax dt [I] where, dP = the number of grams of solute crossing one cm2 in the time dt sec. D, = the diffusion constant for use with fugacities; it is assumed to be constant at a given temperature, and has units of secs. f = the fugacity of the solute in the solid solution; the units are those of pressure, dynes per cm2. x = the distance in cm. Eq 1 can be considered to be the basic form of the first Fick Law for one-dimensional diffusion. In order to convert Eq 1 into an equation involving activities, a choice of a standard or reference fugacity, f°, must be made: then. a = f [2] where a is the activity corresponding to the fugacity f. However, the value of a is also determined by the relation, a = a\c [3] where a is the activity coefficient and c is the concentration. For a given value of a (such as a = 1) it is evident that the units used for expressing concentration will affect the numerical value of a corresponding to a given amount of carbon dissolved in y-iron. Since weight per cent concentration is so widely used, this unit will be adopted as the standard in this paper. When the value off given by Eq 2 is substituted in Eq 1, the first Fick Law becomes, dP= -D,f°dl [4] Since D,f° is constant (at a given temperature), the second form of Fick's Law can be obtained in the usual manner, -dt D'f° dx [5] where c' is the concentration in g per cm3. In terms of weight per cent concentration, c, Eq 5 becomes, dc _ 100 d2a Tt = T /f w [6] where p is the density of the solid solution and will be assumed to be constant. In order to simplify Eq6, it is necessary to consider the standard state, f°. Standard states are chosen for convenience. There appears to be one especially convenient choice for the analysis of interstitial diffusion. If the activity coefficient, a, is set equal to unity at infinite dilution of the solute, 100 then — 100D1fl° is approximately equal P to the usual diffusion constant in dilute solutions. This is true since the activity, a, in Eq 6 can be replaced by the concentration, c, with little error in an infinitely dilute solution.* Here f1° is the standard fugacity necessary to achieve this standard state. It is proposed that a diffusion constant Da' be defined, Da1=100/pD1f1° [7] where A is the metal whose diffusion is * The second derivative of a is equal to the second derivative of c in dilute solution only if -da/dr- is. zero. Although theory does not require this condition, in the systems C in Fe, and Zn in Cu, dc is in fact found to be essentially zero.

Jan 1, 1950

By Shiou-Chuan Sun

THE fact that coals of different ranks and even of the same rank differ greatly in their amenability to iroth flotation is well known. In recognition of the need for an explanation of this phenomenon, two hypotheses have been suggested. Wilkinsl reported that the floatability of coals increased with an increase of the carbon content or rank. This postulate is handicapped by the fact that bituminous coals that possess moderate carbon contents are actually more floatable than anthracite coals that have high carbon contents, as shown in columns 6 and 9 of Table I. Taggart and his associates' implied that the difference of floatability between bituminous and anthracite coal was caused by the variation of carbon-hydrogen ratio. This is not applicable to the relative floatability of other coals and carbons. For example, column 11 of Table I shows that the carbon-hydrogen ratios of low-floating lignitic coal and non-floating animal charcoal are not only smaller than the moderate-floating anthracite coal, but are also similar to the high-floating bituminous coal. Furthermore, according to this hypothesis, high temperature coke-A (464), Ceylon graphite (1238), and lamp-black (357), all possessing extremely high carbon-hydrogen ratios, should be less floatable than other substances having much lower carbon-hydrogen ratios such as high volatile-B bituminous coal (11.9 to 22), anthracite coal (35.7 to 60.5), lignitic coal (15.6 to 33.6), and charcoal (13 to 26.2). However the former group is actually more floatable than the latter group. In this paper, a surface components hypothesis is Proposed to explain the different floatabilities of coals, carbons, and hydrocarbon minerals. The validity of the hypothesis is experimentally supported by the actual floatability, natural floatability, wettability, and adsorbability for neutral oils of coals, carbons, and hydrocarbon minerals tested. The combustible recovery of the flotation results, as used in this paper. was calculated from Eq. 1: P (100-Ep) 100 RWCP Rc= [1] F (100-E,) C, where R, is the percent combustible recovery; F and P are, respectively, the weight of feed and the weight of concentrate or product; E, and Ep are, respectively, the total percent of ash plus moisture in feed and in concentrate; Ru. is the percent weight recovery: and C, and C, are, respectively, the percent of combustible in feed and in concentrate. Except for ash and moisture content, all chemical components of a coal are assumed combustible. The experimental work included studies on flotation, ultimate and proximate analyses, contact angle tests, extractions of bitumen-A with benzene, adsorptions for liquid hydrocarbons, and wetting tests. Most of the flotation experiments were performed in a laboratory Fagergren machine; others were tested in a small Denver machine. The solid feed for the former was 300 g and for the latter was 30 g. The solid materials used for flotation were crushed to —48 mesh. After the mineral pulp in the flotation cell was agitated for 6 min and the pH was adjusted to 7.5 & 0.2 with sodium hydroxide or hydrochloric acid, a petroleum light oil having a viscosity of 5.73 centipoises at 77 °F was added and conditioned for 2 min. Finally, pine oil was introduced and the froth was collected for exactly 3 min. The weight ratio of petroleum light oil to pine oil was kept constant at 1.5 to 1. Tap water was used for all flotation tests. Contact angles were measured with a captive bubble machine. For each coal sample, three specimens were mounted in transoptic mounts and polished with levigated alumina, first on a sheet glass, then on a cloth-covered metal polishing wheel. The polished specimen was first washed with distilled water and wiped thoroughly on a cleaned linen pad, then transferred into the pyrex cell of the captive bubble machine and conditioned for 6 min., and finally measured for contact angles at three or more points. Except where otherwise stated, the induction time for each measurement was 1 min. The contact angle representing each material was obtained by averaging the measurements of three specimens. The linen pad was first washed with warm distilled water, then boiled 30 min in a 2N sodium hydroxide solution, and finally washed with distilled water until no trace of sodium hydroxide could be detected in the decanted solution. The cleaned linen pad was stored under distilled water. Immediately before using, the pad was rewashed and transferred into a clean pyrex petri dish partly filled with distilled water. The glassware and rubber gloves used were cleaned by soaking in sulphuric acid-potassium dichromate cleaning solution, followed by rinsing with distilled water. The polished specimens were handled only by glass forceps. The ultimate and proximate analyses were made in accordance with the ASTM standard procedures for coal and coke. The extractable bitumen-A was determined by weighing 1 g of —100 mesh sample and placing it in a desiccated and weighed ASTM aluminum-extraction thimble. The thimble was placed in condenser hooks and inserted into an extraction flask containing 100 cu cm of benzene. The flask was heated and the benzene vapor was condensed by water coils. At the end of 24 hr of percolation, the thimble was removed, desiccated, and weighed. Loss in weight of sample was taken as bitumen-A and calculated to dry and ash-free basis.

Jan 1, 1955

By Marcus A. Grossmann

The Institute has established this lectureship to honor the memory of a great American metallurgist, one whose fame has continued long after his passing. As one scientist recently stated it, "All metallurgical texts gradually become outdated and outmoded, except those of the great Henry Marion Howe." It seems fitting, then, that, once every year, we shouId review, in his honor, some aspects of the science in which he was such a distinguished leader, and also discuss any new data that may help to clarify our concepts. It is my privilege, for which indeed I express the deepest appreciation to this Institute, to offer for consideration a few thoughts on the toughness and fracture of hardened steels. Before proceeding to the data, I wish to express my great indebtedness to many associates whose stimulating discussions, generous contributions of data and cooperation in experiments are gratefully acknowledged. (But I wish to absolve them of any blame that may arise from the way in which their discussions and data have been used!) First, Dr. Edgar C. Bain, Vice President, Carnegie-Illinois Steel Corporation, whose amiable but merciless criticisms and always constructive and thoughtful suggestions help unfailingly to guard against pitfalls. Then to associates at Gary Steel Works, especially Mr. H. B. Wishart, Mr. W. K. Smith and Mr. V. Elliott, for great assistance in securing data. Also to Col. C. H. Greenall, Director of Laboratories, Frankford Arsenal, and his metallurgical staff, including Messrs. J. K. Desmond, H. Markus, D. F. Armiento, H. Rosenthal and Henry George, for numerous data and illustrations. Further to Dr. J. R. Low and Dr. M. Asimow, for helpful discussions of effects of states of stress during deformation and rupture, to Mr. J. M. Hodge for discussions and data, to the staff of the U. S. Steel Corporation Research Laboratory for advice and data, and to the Metallurgical Staff at Duquesne Steel Works. The individual acknowledgments will also be found under illustrations. Acknowledgment is also due to Mr. S. C. Snyder for incredibly meticulous scrutiny of manuscript. To all of these, grateful thanks. Introduction The discussion to be offered here will deal with quench-hardened steels only, and may perhaps be considered under three headings: 1. It was found that, when ferrite was precipitated in the (prior austenite) grain boundaries of hardened steel, the notch-bar toughness was much less than in pieces fully hardened. The fracture took place in the ferrite, and the loss in toughness is attributed to the lower cohesive (cleavage) strength of the ferrite, as cornpared with martensite. 2. When notch-bar impact test pieces of quenched and tempered steels were examined as to their mode of fracture, it

Jan 1, 1947