Search Documents

By Earle Faig

Earle Faig Along with other transportation modes, the barge industry will be greatly affected by trends in domestic coal production. Traditionally, coal has been one of the major commodities moved by barge. Now, because of events in the international energy market, coal is finally getting the recognition it deserves as a major component of the country's energy program. Perhaps even more significantly, a series of events has caused other countries to turn to the United States as a source of coal. Although the outlook for coal demand is optimistic, it will be greatly influenced by the ability of the transportation sector to handle the additional traffic. Because of its long experience in moving large quantities of coal, the barging industry is in an excellent position to capitalize on coal's increased popularity. A brief description of the barging industry should facilitate the discussion of coal as it relates to that industry. Barge companies operate on a river system that is over 40 Mm (25,000 miles) long. It includes the East and West Coasts and the Mississippi River System comprised by the Allegheny, Monongahela, Kanawha, Ohio, Illinois, Missouri, Tennessee, Cumberland, Arkansas, Warrior-Tombigbee, and Mobile Rivers, and some smaller tributaries. The Gulf Intracoastal Waterway stretching along the Gulf Coast from Brownsville, TX, to Apalachicola, FL, also is considered part of the inland waterway system. Approximately 1,800 companies operate on the inland river system and have a combined fleet of about 4,400 towboats and 28,000 barges. The fleet of towboats represents 4.8 GW (6.4 million hp) of power. The towboats comprising this fleet vary in size from a small harbor boat to a large line haul boat with 7.5 MW (10,000 hp) of power. The choice of size and power of the boat depends on a number of factors, including the size of the river and operational constraints, such as lock size, river-current rate, and water level. The choice also depends on the tow size. A single tow on a river capable of locking through 15 barges at a time handles up to almost 20 kt (22,000 st), whereas a tow on the lower Mississippi, which has no locks, can accommodate up to 41 kt (45,000 st) of cargo. Because the quality of coal is not typically damaged by exposure to the elements, it is usually moved in open hopper barges, although covered barges are used occasionally. The most frequently used size is the jumbo barge that measures 59 m (195 ft) long by 11 m (35 ft) wide by 4 m (12 ft) deep and can handle up to 1.4 kt (1,500 st) of coal. Advantages of Barge Transportation Barging offers several inherent advantages in transportation of bulk commodities like coal. First, barges can carry large quantities of product. A single jumbo hopper barge can handle 1.4 kt (1,500 st) of cargo while it takes 15 91-t (100-st) capacity rail cars or 57 trucks hauling 24 t (26 st) each to move an equivalent amount of cargo. Thus, barges can move more tonnage with less equipment than competing forms of transportation. Second, barging offers the most fuel efficient means of transportation. Less energy is required to push a load over water than to pull a similar load over land. The amount of pollution attributable to barge traffic is minimal because barges consume less fuel. Third, barging is an inexpensive form of transportation. The inland waterway industry transports about 12% of the nation's freight for about 2% of the total freight bill. This can be attributed to competition between the large number of companies on the river system and competition with other transportation modes. Fourth, barging is an environmentally safe method of transportation. Inland river routes expose fewer urban areas to hazardous substances than rail or truck. And barge spills occur less frequently than rail or truck spills. The industry does pose some disadvantages. The waterway system is constrained by the natural river network, so barges cannot travel on the most direct straight-line route. This means that barging is usually an incomplete form of transportation requiring interface with truck or rail. The industry's market is limited to those geographical areas that are located on or near the rivers. In addition, it is a time-consuming transportation mode, making it impractical for many types of commodities.

Jan 10, 1982

The Annual Meeting of The Rocky Mountain Coal Mining Institute was called to order by the President, John T. Atkins, at 2:00 p.m. Tuesday, June 26, 1984 in the Colorado Ballroom of the Marriott's Mark Resort in Vail, Colorado. There were 36 members in attendance. It was moved by Bob Diederich, seconded and carried to dispense with the reading of the 1983 Annual Meeting minutes as printed on pages 105 thru 109 of the 1983 Proceedings. In keeping with the printed agenda in the convention program/ticket booklet, the President called for reports of standing committees as follow: 1. Bylaws & Resolutions - Loren Linville, Chairman Linville noted that the committee had been working for two years on bylaws changes and concluded a corporate lawyer specializing in nonprofit organizations be engaged. He then introduced Mr. Peter B. Nagle of the Denver law firm of Sherman and Howard who outlined the major changes in converting the Institute's current status of 501(c)(6) to 501(c)(3). He pointed out that such a conversion would give the Institute a proper legal status, would give individual contributors to the scholarship or other programs a personal income tax deduction, would qualify the Institute for the postal nonprofit mailing rate and the Institute would achieve the advantage of limited liability inherent in a corportion and would insulate its members, officers and directors from liability for corporate obligations. Copies of the Articles of Incorporation as well as the Bylaws were available for all members. A discussion ensued and the following resolution was moved by Bob Diederich, seconded and carried: RESOLVED, that the unincorporated association known as the Rocky Mountain Coal Mining Institute be, and hereby is, dissolved and in its place, and to carry out its objectives, shall be formed a nonprofit corporation under the laws of the State of Colorado, which corporation shall also be known as The Rocky Mountain Coal Mining Institute, and to which shall be transferred all of the assets and liabilities of this association. FURTHER RESOLVED, that the proper officers of this association are authorized and directed to execute and file with the Colorado Secretary of State, Articles of Incorporation in a form substantially similar to those prepared for this association by its counsel, and the names of the individuals identified in such Articles of Incorporation as the initial Board of Directors of such corporation, which names were proposed and recommended by the Nominating Committee of this association, are hereby approved. FURTHER RESOLVED, that the proper officers of this association are authorized and directed to do all acts necessary or convenient to carry out the terms of the preceding resolutions. (Articles of Incorporation and Bylaws as adopted are printed following these minutes.) Additional resolutions were read by Linville as follows: Item 1: WE RESOLVE herein that in the year since the last convention of the RMCMI, the Institute and the coal industry have suffered the loss, through death, and the services of: Lloyd W. Taylor Charles T. Holland THEREFORE, BE IT RESOLVED, that we record in the minutes of the RMCMI, recognition of the passing of these members who have so greatly contributed to the advancement of the industry, both nationally and in the Rocky Mountain area, and that a copy of this resolution be sent to the next of kin as evidence of the sincere sympathy and deep regard of the members of the Institute. Item 2: RESOLVED, that in recognition of their long and faithful service to the coal industry that the status of Life Member be conferred upon: Bernard Bernstrom Roy Coulson Shirl C. McArthur Item 3: RESOLVED, that we commend and express our gratitude to our Executive Secretary, officers, chairmen and committees for an outstanding job in presenting. the 80th RMCMI meeting. Item 4: RESOLVED, that the RMCMI express its gratitude to K&P, INC. of Golden, Colorado for their generous donation of office facilities and furniture for the RMCMI office and to K&P SALES ENGINEERS, INC. also of Golden, Colorado for their assistance in telephone answering, other office procedures and the use of certain office equipment. Item 5: RESOLVED, that six scholarships be awarded annually to advance the education program of the Institute to worthy students pursuing career paths in mining or other related industry in the six states in which the Institute generally operates and has members and in accordance with the criteria set forth by the Scholarship Committee. It was moved by Jim Neal, seconded and tamed

Jan 1, 1984

By R. P. Abendroth

This study is concerned with the kinetics of selective oxidation of chromium in a commercial Fe-Cr-Ni alloy. Selective oxidation of chromium in this alloy, by use of a low oxygen-potential atmosphere, leads to the formation of a compact, protective layer of Cr2O3. This layer serves to protect this alloy from gross scaling when it is subsequently exposed to severe oxidizing conditions at high temperature. The interaction of low oxygen-potential atmospheres close to equilibrium with Fe-Cr and Ni-Cr alloys has been studied by others."' These studies werk concerned with the surface-structure variations under slightly oxidizing conditions. NO detailed study was made of the oxide scale formation kinetics, however. The alloy samples were cut from 0.012-in.-thick sheet, with an apparent surface area of 7.1 sq cm. These sheet samples were abraded through 4/0 metallographic paper, and washed in alcohol and acetone. The analysis of the sheet alloy is (in weight percent): 42 pct Ni, 5.5 pct Cr, 0.09 pct C, 0.18 pct Al, 0.36 pct Si, 0.26 pct Mn, balance Fe. The weight gain vs time data were obtained with a 2-g capacity fused-silica spring—cathetometer system. The spring deflection was optically magnified ten times before being read by the cathetometer. A sensitivity of about 0.01 mg was attainable. The spring was enclosed in a water jacket maintained at 60°C to minimize the effect of temperature changes. The sample was suspended from the spring with a fused-silica hangdown and was positioned in the thermal center of a mullite furnace tube. Sample temperature was read with the aid of a thermocouple placed outside the mullite tube and an inside vs outside temperature calibration. Temperature change during the course of a run was ±0.25°C, with a temperature gradient of less than l.O°C over the length of the sample. Total temperature uncertainty was no more than ±3.0°C. Alignment difficulties between the hangdown and radiation shields in the furnace tube required that the sample be positioned in the furnace when cold, and heated with the furnace until the temperature stabilized at the desired point. This required 5 to 6 hr, and was carried out in Matheson ultrahigh-purity hydrogen. Oxidation was started, after evacuation, by introducing a hydrogen-water vapor mixture, obtained by saturating hydrogen with water vapor at 31.00o ± 0.02oC. Oxidation was continued for 90 min. Gas flow was 300 ml per min during heat up and oxidation. Since only several milligrams of oxide are formed on each sample, chemical analysis of the oxide is difficult. A representative analysis is: 80 pct Cr2O3, 5 pct Fe2O3, 3 pct Al2O3, 4 pct MnO, 7 pct SiO2, 1 pct or less NiO. X-ray diffraction analysis of the oxide as formed on the alloy gives rhombohedra1 Cr2O3, and barely distinguishable amounts of a cubic spinel phase, and possibly AlZOs and SO2. The identification of these latter two compounds is by no means certain. The spinel phase could be based on iron or manganese as these elements are present in significant amounts in the oxide. The results of the kinetic studies using 31°C dew-point hydrogen-water vapor mixtures were found to conform to a parabolic rate law. In many cases the parabolic plot consisted of two intersecting straight lines, defining an early and a late rate for a particular run, and in the other cases the parabolic plot consisted of one straight line for the entire run. The slopes of the various straight lines were determined by the method of least squares. Reproducibility of the data was good enough for multiple runs at the same temperature such that the value given for the rate constant is the average for two or more closely similar values, rather than widely varying values of the rate, where more than one determination is indicated in Table I. The exhibition of only one or of two rate constants during a run can happen at the same temperature. Thus, Table I shows that at 11'74°C a single rate constant was obtained from one sample, while other samples oxidized at the same temperature gave an early and a late rate constant. It should be noted that the single rate constant corresponds very closely with the early rate constant. This is also true at 1153°C. The time at which the late rate started to appear was variable, usually occurring 20 to 40 min after oxidation had started.

Jan 1, 1964

By S. H. Williston

IT not often that bore hole surveys can be checked by actual. civil engineering methods. A recent Arizona survey was checked by normal surveying methods and the comparison of the results should be of value to both oil and mining men. During the summer of 1948 the Phelps-Dodge Corporation, at its Copper Queen property near Bis-bee, Ariz., drilled a 1245 ft, 8 in. diarn, churn drillhole in a mineralized area and cased part of it, intending to use it to transfer mill tailings for stope fill. The hole, as frequently occurs, was not straight, and, in endeavoring to locate the bottom in the underground workings, they found no evidence of the hole at the underground coordinates directly below the surface location. The noise of the drilling tools was reasonably clear, but the direction of sound was uncertain. Preliminary tests with available equipment were not successful in locating the bottom of the hole. Because of the mineralized character of the area and the fact there was some casing in the hole, any magnetic method of well surveying would give results of doubtful value. Sperry-Sun gyroscopic well-surveying instruments were finally used to locate the bottom of the hole. These instruments consist of a gyroscope to determine azimuth and either a pendulum or bubble inclino-meter. A multiple shot camera photographs both instruments on a single film and superimposes the photograph of a watch. Coordination of depth with time at the surface makes it possible to select the corresponding picture for any depth. After making several runs of the empty instrument housing from the top of the hole to the bottom to make sure there were no obstructions in the hole, three surveys on wire line were completed during the afternoon. The three surveys, in which readings were taken at different points in the hole on each survey, were computed and gave the following locations of the bottom of the hole in relation to the surface collar: survey No. 1—24.92 N, 30.30 W; survey No. 2—24.24 N, 31.11 W; survey No. 3—26.54 N, 27.72 W. Then the data from the three surveys were combined into a single set of calculations which gave a location for the bottom of the hole: combined surveys—24.27 N, 30.16 W. (Fig. 1.) Immediately upon the determination of the coordinates at the bottom of the hole, a drift on the 1300 ft level was started toward the indicated loca- tion some 38 ft northwest of the coordinates of the surface location, The bottom of the hole was located within the drift round in which it was expected, and the transit survey run to the actual location of the hole indicated N 27.18, W 29.71. This shows a discrepancy between the well survey and the transit survey of 0.45 ft in the westerly direction and 2.91 ft in the northerly direction. All surveys, both gyroscopic and transit, fell well within the width of an ordinary drift. While this is satisfactory for almost any and all mining requirements, a theoretical examination was made as to reasons for the discrepancy. A study of the course of the hole indicates that considerable right turn or spiral existed, and in all probability the surveying instrument was pulled out of alignment while traversing the turn by approximately 0.05 ft at the top and another 0.05 ft in the opposite direction at the bottom of the instrument. If such an allowance were to be made, the survey calculations would almost exactly correspond with those determined by transit. This sort of discrepancy would be minimized by the use of stabilizing guides. It is unfortunate that physical laws probably effectively prevent the use of gyroscopic instruments in EX and AX diamond drill holes. The directive power of a gyroscope falls off inversely at some rate between the third power and the sixth power of the diameter. Present instruments can be run in casing 53/4 in. ID or over and might be adapted to somewhat smaller diameters, but the difficulty of reducing these diameters to 11/4 in. or 2 in. is almost insurmountable at the present time. Acknowledgment The author wishes to express his appreciation to the Phelps-Dodge Corporation for permission to publish this article, and to the Operating and Engineering Departments for their cooperation on the survey; also to Donald Hering, of the Sperry-Sun Well Surveying Co., who actually made the survey and calculated the results.

Jan 1, 1951

By S. H. Williston

IT not often that bore hole surveys can be checked by actual. civil engineering methods. A recent Arizona survey was checked by normal surveying methods and the comparison of the results should be of value to both oil and mining men. During the summer of 1948 the Phelps-Dodge Corporation, at its Copper Queen property near Bis-bee, Ariz., drilled a 1245 ft, 8 in. diarn, churn drillhole in a mineralized area and cased part of it, intending to use it to transfer mill tailings for stope fill. The hole, as frequently occurs, was not straight, and, in endeavoring to locate the bottom in the underground workings, they found no evidence of the hole at the underground coordinates directly below the surface location. The noise of the drilling tools was reasonably clear, but the direction of sound was uncertain. Preliminary tests with available equipment were not successful in locating the bottom of the hole. Because of the mineralized character of the area and the fact there was some casing in the hole, any magnetic method of well surveying would give results of doubtful value. Sperry-Sun gyroscopic well-surveying instruments were finally used to locate the bottom of the hole. These instruments consist of a gyroscope to determine azimuth and either a pendulum or bubble inclino-meter. A multiple shot camera photographs both instruments on a single film and superimposes the photograph of a watch. Coordination of depth with time at the surface makes it possible to select the corresponding picture for any depth. After making several runs of the empty instrument housing from the top of the hole to the bottom to make sure there were no obstructions in the hole, three surveys on wire line were completed during the afternoon. The three surveys, in which readings were taken at different points in the hole on each survey, were computed and gave the following locations of the bottom of the hole in relation to the surface collar: survey No. 1—24.92 N, 30.30 W; survey No. 2—24.24 N, 31.11 W; survey No. 3—26.54 N, 27.72 W. Then the data from the three surveys were combined into a single set of calculations which gave a location for the bottom of the hole: combined surveys—24.27 N, 30.16 W. (Fig. 1.) Immediately upon the determination of the coordinates at the bottom of the hole, a drift on the 1300 ft level was started toward the indicated loca- tion some 38 ft northwest of the coordinates of the surface location, The bottom of the hole was located within the drift round in which it was expected, and the transit survey run to the actual location of the hole indicated N 27.18, W 29.71. This shows a discrepancy between the well survey and the transit survey of 0.45 ft in the westerly direction and 2.91 ft in the northerly direction. All surveys, both gyroscopic and transit, fell well within the width of an ordinary drift. While this is satisfactory for almost any and all mining requirements, a theoretical examination was made as to reasons for the discrepancy. A study of the course of the hole indicates that considerable right turn or spiral existed, and in all probability the surveying instrument was pulled out of alignment while traversing the turn by approximately 0.05 ft at the top and another 0.05 ft in the opposite direction at the bottom of the instrument. If such an allowance were to be made, the survey calculations would almost exactly correspond with those determined by transit. This sort of discrepancy would be minimized by the use of stabilizing guides. It is unfortunate that physical laws probably effectively prevent the use of gyroscopic instruments in EX and AX diamond drill holes. The directive power of a gyroscope falls off inversely at some rate between the third power and the sixth power of the diameter. Present instruments can be run in casing 53/4 in. ID or over and might be adapted to somewhat smaller diameters, but the difficulty of reducing these diameters to 11/4 in. or 2 in. is almost insurmountable at the present time. Acknowledgment The author wishes to express his appreciation to the Phelps-Dodge Corporation for permission to publish this article, and to the Operating and Engineering Departments for their cooperation on the survey; also to Donald Hering, of the Sperry-Sun Well Surveying Co., who actually made the survey and calculated the results.

Jan 1, 1951

By Dilip K. Das, Douglas T. Pitman

ONE of the major uses of barium in metallic form is as a getter material in vacuum tubes. Because of the high chemical reactivity of the metal, Ba-Al alloys are extensively used. Numerous methods for the preparation of Ba-Al alloys have been published, a few of which1-4 are cited here. Most of these methods were found to be quite elaborate, involving the reduction of BaO, and not too well adapted for the close control of the final composition of small amounts of alloys prepared for laboratory use. A simple laboratory method for the preparation of Ba-A1 alloys in small batches starting from pure metals was devised, so that it was possible to control the desired compositions to within 1 pct. The pertinent features of the alloy system Ba-Al5 are 1) an intermediate compound BaAl, with the melting point of 1050°C, and 2) a eutectic between aluminum and BaAl4 at 98 pct Al. The accompanying sketch shows the experimental arrangement for the preparation of the alloys. Weighed amounts of aluminum and barium were placed in an alumina and a stainless steel crucible, respectively. According to the supplier's specification, the purity of the metals used in the alloys is as follows: a) aluminum rods—99.9 pct Al, and b) barium rods—99.5 pct Ba. The stainless steel crucible, tapered at the bottom and having a 1/16 in. diam hole, rested on top of the alumina crucible. The assembly was placed inside a graphite sleeve which rested on a refractory platform. The platform moved the assembly up and down through the field of a radio frequency coil. A glass bell jar was placed between the crucible assembly and the radio frequency coil to maintain a steady flow of helium around the melt. A small window was cut out on the wall of the alumina crucible to observe the progress of the re- action and to record the temperature with an optical pyrometer. The platform was first raised high enough to move the barium out of the radio frequency coil field in order to allow only the aluminum to melt. The assembly was then lowered so that the barium began to melt and flow out through the small orifice into the molten aluminum. In order to keep the violence of the exothermic reaction under control, the rate of flow of barium was carefully regulated by raising or lowering the crucible assembly. All the samples prepared by this technique were examined by a Norelco X-ray diffractometer using CuKa radiation, The diffraction specimens were prepared by placing the finely powdered samples in flat specimen holders. The Ba-Al alloys prepared with a high barium content were found to consist mainly of BaAl4. The structure of BaAl4 has previously been reported by Alberti and Andress.8 They found that BaAl4 was body-centered-tetragonal with an a0 = b0 = 4.530Å and c0 = 11.14Å. An alloy whose composition was found by chemical analysis to be almost 100 pct BaA1, was used to determine the relative intensities. The d-spacings were obtained from the same alloy to which a small amount of tungsten had been added as a calibrating material. Accurate values for a, and c, were calculated according to the method proposed by Taylor and Floyd.' The calculated values are: a, = b, = 4.566Å and c0 = 11.250Å. The measured d-values for BaAl4 are shown in Table I along with relative peak intensities above background and hkl indices. Acknowledgment The authors are grateful to L. J. Cronin, the head of the Techniques Dept., for suggesting the problem and for his constant interest. References 1Froges and Camargue: German Patent No. 809107, 1951. French Patent No. 935324, 1949. 2E. Bonnier: Annales de physique, 1953, vol. 8, pp. 259-312. 3M. Orman and E. Zemhela: Prac Institute of Metals; 1952, vol. 4, pp. 437-445. 4E. Fujita and H. Yokomizo: Reports Gov. Chemical Industrial Research Institute, Tokyo, 1952, vol. 47, pp. 291-297. 5E. Alberti: Ztsch. fur Metallkunde, 1934, vol. 26, p. 6. 6E Alberti and K. R. Andress: Ztsch. fur Metallkunde, 1935, vol. 27, p. 126. 7 A. Tnslor and R. W. Floyd: Acta Crystallographica, 1950, vol. 3, p. 285.

Jan 1, 1958

By W. E. Rudolph, R. E. Baylor

DUE to its location in the Atacama Desert, one of the most barren of the earth's surfaces, Chuquicamata's water supply presents unusual problems. Yearly rain-fall averages less than one tenth of an inch at the plant. However, there are summer showers above 12,000 ft in the Cordillera to the east, the resulting run-off flowing through old river valleys buried beneath more recent volcanic formations, to be impounded within sediment-filled basins. This water emerges at springs where the outlets of these basins are blocked by lava flows, and here are formed the small streams which feed the only important river of the region, the Rio Loa. Chuquicamata's water is obtained from these springs and rivulets. [ ] The map above indicates four pipe lines from which potable and industrial water are supplied. Potable water, amounting to 4500 metric tons per day, is conveyed in the Toconce pipe line from springs 59 miles due east of Chuquicamata. This water is used not only for drinking, but also for boilers and other needs requiring high quality. For industrial water at the oxide plant, there are two 12-in. pipe lines from the Rio San Pedro, carrying a total of 17,000 metric tons per day of slightly brackish water. This water is at present used mainly for leaching and for hygienic purposes. Water Source Found For the present and future needs of the sulphide plant, it was calculated that at least 32,000 metric tons per day of make-up water would be required. For this purpose, a pipe line of 44 miles length was constructed to bring in the entire flow of the Arroyo Salado, one of the eastern tributaries of the Loa. The salt content of this water is so high (over 5000 parts per million of solubles, mostly chlorides) that it is highly detrimental to farming, and the Chilean Government had been studying projects to separate these waters from others of the Loa system in order to improve agricultural conditions in the fertile valley of Calama. So it happened that the Government was willing to award rights to the Arroyo Salado waters under agreement whereby the Mining Company removes waters from the Rio Loa system above Calama for all time. The outlet of these waters, after serving their purpose at the new concentrator and leaving the plant in tailing, is the Salar de Talabre, an old salt lake which presents fully ten square miles of surface to serve as an evaporating pan, the outlets having now been blocked by dams. Here the dry climate of Chuquicamata is a favourable factor, evaporation averaging slightly above 1/4 in. per day. The Toconce and San Pedro pipe lines have been functioning from 26 to 34 years, and through the use of special cleaning tools which were developed at the plant, as well as deaeration of the more active Toconce water, these pipes are now maintained at capacities which do not diminish as years go on. Constructing the Dam The Arroyo Salado pipe line design and construction involved certain special and interesting features, and inasmuch as this line and its intake works are solely for the needs of the new sulphide plant, more detailed description is given. The waters are impounded at a gravity dam constructed of concrete to a height of 100 ft above the river bed, keyed into the precipitous Dacite walls of the narrow canyon (barely 6-ft wide at the bottom, only 25-ft width at 50 ft above). A small secondary dam was built 100 ft down stream from the main dam, providing a pool of 15-ft depth to protect the main structure from flood flows over the spill-way during the rainy season. A system of four 36-in. syphons was designed for discharging these flood waters from the lower depths of the lake, in order to avoid eventual sedimentation behind the dam. The lake has a length of 3300 ft, and its water level is controlled by an adjustable spillway permitting draw-down of eighty inches, amounting to 41,000 metric tons of available capacity. This regulation is necessary because of wide fluctuations in stream flow between day and night due to freezing of feeders. During the construction of the dam the entire river flow was handled within a 36-in. pipe line some 2000 ft in length. As the excavations proceeded

Jan 1, 1952

By W. F. Riggs

INTRODUCTION Flotation equipment is becoming more definitive in design as well as greater in dimension. The last ten years have witnessed the implementation of large flotation cells ranging in size from 28 to 85 cubic meters. However, small plants still operate with cells as small as 3 cubic meters. These changes have increased the operational sensitivity of the flotation system to the chemical reagents used. No scientific study has been undertaken to develop the relationship between the size of equipment and the response of reagents in the flotation system. Hence, this article is an attempt to present several observations from past experience for the purpose of stimulating future research in an area previously ignored and perhaps subsequently not used to advantage. CONVENTIONAL FLOTATION Recent advances in conventional flotation cell design have been made in two areas: a. Flotation cells are becoming larger, and b. Basic design of flotation cells has changed. They can be classified into two types: deep cell and shallow draft cell. Size and design changes of flotation cells have a greater effect on frother performance than on the performance of other chemical reagents. Hence, the discussion in the following sections is restricted to the impact of frothers in the flotation system. Flotation Cell Size The advent of increasing flotation cell size has resulted in the increase of frother consumption. There appears to be a correlation between reduced horsepower (kilowatt) consumption and increased frother consumption per ton of ore processed in the circuit. This is more prevalent when the frother chemistry remains unchanged. Another reason for increased frother consumption is the increased distance between the rotating central shaft and the lip of the cell for larger flotation machines. Increasing this distance results in increased surface area corresponding to the larger cell size. Reduced power input per ton combined with the need for longer, stronger load support of the froth bubble as a result of increasing cell size has resulted in the development of a new series of stronger and more persistent frothers. By the same token, these newer, stronger frothers are difficult to control when applied to the smaller cells. This is true even at lower dosages because the break point between too much frother and not enough frother for optimum metallurgy is normally very narrow. This problem is often exaggerated when operators have established an optimum setting for pulp level and air for the particular frother in use. Generally, the tendency is to operate the flotation cells simply by adjusting frother dosage. The increase in frother strength to accommodate the large cells often results in the realization, after a retrofit, that the existing pump box or sump which is still handling the same volume of concentrate is no longer adequate in size. Flotation Cell Design The two most common flotation cell designs installed during the past ten years are the deep cell and the shallow draft cell. The deep cell provides for complete mixing of air and pulp at the bottom of the cell. The shallow draft cell does all the mixing/pumping in the bottom haft of the cell whereas the mixing of the air is performed further up the shaft and is brought into contact with the pulp midway up the vortex. There are many preferences for one or the other. It is the author's contention that both designs have very valid concepts and must simply be suited to the particular situation as required by the operator. The following comments are subjective observations which should be explored in greater detail. The deep cell mixes both air and pulp together at the bottom of the cell requiring a full depth ascending of the homogenized air to the surface to form the froth layer. This design, in combination with charged air, provides for good concentrate grade control under adverse conditions of selectivity, especially when faced with a fine grind and a slime-laden gangue. A selective frother can be applied without a competitive loss in recovery. One of the conditions for good selectivity in this environment is operating at medium pulp solids concentration in the range of twenty-six to thirty-four per cent (26-34%). When a higher solids concentration is used in combination with a coarse grind, it becomes difficult to maintain the same recovery with a selective frother due to higher pulp viscosity. This condition sometimes results in an unacceptable coalescence of the air bubbles as they ascend the full height of the deep cell. The increased pulp viscosity at higher solids concentration can also retard the flotation kinetics. The combination of decreased surface area of dispersed air and slower kinetics can result in reduced recovery causing coarse particles to continue to drop

Jan 1, 1989

By W. J. Schuster

Safety in coal mines depends largely upon adequate training of the foreman. Although management must provide modern and safe equipment and at all times keep mines in first class condition from a safety viewpoint, final results will be determined by the quality of supervision. HANNA COAL CO., Division of Pittsburgh Consolidation Coal Co., operates three large underground mines in eastern Ohio. The section of Pittsburgh No. 8 coal seam in which these mines are located varies in thickness from 52 to 64 in. It is immediately overlain by a stratum of shaly material 12 to 15 in. thick locally known as draw slate, which is structurally very weak and which disintegrates rapidly upon exposure to atmosphere. Immediately above the draw slate as it is normally found is a band of extremely high ash material 6 to 12 in. thick known as roof coal or rooster coal, and above this is a stratum of conglomerate material varying from 4 to 10 ft in depth. Overlying the conglomerate is a relatively thick stratum of limestone, the first stable material above the Pittsburgh coal seam in eastern Ohio. With the method of full-seam mining that has been adopted, draw slate is shot down, loaded with the coal, and removed in the preparation plants. The roof coal then becomes the permanent roof. The major problem in mining the No. 8 seam in eastern Ohio is control of the roof. Since the strata above the draw slate contains no material with a structure firm enough to provide self-support, the roof begins to sag in a relatively short time after the coal and draw slate have been removed. The problem thus becomes one of getting temporary safety posts under this roof as quickly as possible to prevent a break or separation from occurring either in the roof coal or in the conglomerate above it. Haulage System The Pittsburgh No. 8 seam in eastern Ohio is relatively level, with only minor local dips. Throughout the Hanna Coal Co. mines, entries are generally 12 ft wide. Rooms are driven on a 60" angle on 30-ft centers and are 22 ft wide. No attempt is made to extract the 8-ft pillars between. The entire length of main line haulage is gunited in one mine, and a major portion in another. Two of the mines have single-track main haulage roads with passways. The third, a new mine, is double-tracked, and the roof is supported by steel crossbars, 60 lb or heavier, spaced on 4-ft centers and lagged. In recent years timbering on main line and secondary haulage roads has been accomplished by one of two methods: 1—crossbars are supported on a small section of post set in a hitch hole in the rib, or 2—or a hole is drilled in the rib about 12 in. below the roof, of sufficient depth to fasten securely a short length of 40-lb rail, the bottom of the rail facing the roof, on which a short post is set directly under the crossbar. At present the hitch-hole timbering method is favored. At two of the mines the main line haulage locomotives are 26-ton, 8-wheel units. These locomotives are of the axleless type, each wheel being individually mounted on the frame. The motorman's compartment is encircled by 3-in. armor plate for the protection of the occupants. At the third underground mine conventional 15-ton locomotives are being used. However, these locomotives have been completely rebuilt in the company's shops. Equipment has been streamlined and quarters have been provided for two people, who are protected by heavy steel plate in much the same way described above. This modernization program has been completed on all secondary haulage locomotives at the three mines, and the company is well on the way to similar equipment of the 6-ton section locomotives. The following additional features have been included in their modernization: 1—additional support for the motors to prevent their falling to the middle of the track and derailing the locomotives should a break occur in the suspension bar support; 2—installation of additional bracing to prevent brake rigging from becoming displaced and causing derailments; 3—enclosure of all electric wiring in conduit or raceway; 4—provision of an enclosed compartment for the storage of re-railers, jacks, and other equipment, so that they need not be carried on the outside of the motor; and 5—redesign of the end of the locomotive opposite the operator's compartment to prevent anyone's mounting from that direction. It is interesting to note that some

Jan 1, 1955

By G. B. Misra

While investigating on the aerodynamic resistance of airways supported with peripheral timber sets, at regular intervals, the following theoretical equations were developed by the author to estimate the resistance coefficient of such airways: [ ] for S < 1, where f is Darcy-Weisbach resistance coefficient of the airway, C is modified drag coefficient of the supporting member, D is equivalent diameter of the bare airway, 8 is ratio of the approach velocity over the sets to the average velocity of the bare airway, A is cross-sectional area of the bare airway, a is projected frontal area of the sets, A., is cross-sectional area of the air stream at the vena contracta inside the set, S is spacing of the sets, f, is resistance coefficient of the bare airway, l is length of aerodynamic influence of sets, p is perimeter of the bare airway, p, is setted portion of the perimeter of the bare airway, pe is unsetted portion of the perimeter of the bare airway, and P shielding factor. The equations were verified experimentally in a model rectangular airway supported with one- (bars), three-, and four-piece sets of square-section timber of three different sizes and were found to hold true. The work has been further extended to one-, three-, and four-piece sets of round timber of 2.6, 3.2, and 3.8 cm diam with the same experimental set up. Tests have been carried out for spacings of 25, 50, 75, 100, 150, and 200 cm over a regime of flow defined by the Reynolds number (with respect to the equivalent diameter of the bare duct) ranging from about 1.5 X 106 to 5 X 106 using the same experimental techniques. The values of f are calculated in the manner indicated in [Ref. 1]. Unlike with square-section timber, the resistance coefficient f of the airway setted with round timber shows a distinct variation with the Reynolds number of flow. This conforms to observations made by Sales and Hinsley.2 In order to have a comparable value of f for all types of sets with all sizes of timber, it was necessary to select the value of f at a fixed Reynolds number of flow. Since f is chiefly a function of the drag coefficient of the sets, the appropriate Reynolds number RE is that with respect to the diameter of timber in the set. Considering the diameters of timber used and the regime of flow over which measurements were made, f was chosen at a value of RE = 20,000 in all cases. The f vs. S curves are maximal in nature and in conformity with theory, the f vs. 1/S curves are straight lines up to a value of S = 1 beyond which they show a distinct flexure. The observed values of 1, the length of aerodynamic influence of sets, agree with the relation 1 = 42 e, developed for square-section timber sets, thus suggesting that the shape of timber has little influence on the length of aerodynamic influence. The value of the modified drag coefficient CD for round timber was calculated in the same way as for square timber in Ref. 1, taking the contraction factor Z = 1.5 for round-edged constrictions. CD has an average value of 0.96 with a standard deviation of 6.08% as compared to the free stream drag coefficient of 1.2 at RE = 20,000 for long cylindrical obstructions The shielding factor [ ] is plotted against S/1 in [Fig. 1]. The curves are more or less independent of the size of timber, but are different for the different types of sets, possibly due to their different degree of symmetry. Values of f calculated by the author&apos;s [Eqs. 1 and 2], using experimental values of CD&apos; and [ ] and taking I = 42 e, are plotted in [Fig. 2] against experimentally measured values of f for different types of sets with different sizes of round timber. The values agree closely with a standard deviation of only 5%, thus establishing the veracity of the theoretical equations developed by the author for round timber as well. A comparison was made between the Xenofontowa4 equations (the only other reasonable relations available for the estimation of the resistance coefficient of supported airways) and the author&apos;s [Eqs. 1 and 2] by comparing in [Fig. 3] the values of the resistance coefficient f computed by the Xenofontowa relations with those experimentally measured by the author. In order to make

Jan 1, 1975

By C. S. Barrett, A. Smigelskas

There have been no publications of the deformation and recrystallization orientations of the metal beryllium, yet pronounced textures would certainly be anticipated since it is close-packed hexagonal in structure. Having an axial ratio approximately that of magnesium, beryllium probably deforms by nearly the same slip and twinning mechanisms that operate in magnesium, and the textures are likely to be similar or but slightly different from the magnesium textures. In the tests reported below this is found to be the case; the textures are found to differ from those of magnesium only in the details of the scatter from the average orientation. This report covers not only samples rolled at room temperature, but some rolled at elevated temperatures. Since magnesium has been suspected by some investigators of altering its crystallo-graphic deformation mechanism at elevated temperatures, it was considered possible that beryllium might do so and alter its textures accordingly. No pronounced alterations were found, however. Unfortunately, the theory of deformation textures is not in a state of development that permits one to deduce the deformation mechanism from a knowledge of the textures, which means that the similarity of textures at different rolling temperatures, reported here, cannot be taken as definite evidence that the deformation mechanism is actually the same at all temperatures. The general similarity of the deformation textures of magnesium and beryllium also extend to the recrystallization textures of the two metals, judging by the pole figures for recrystallized sheet presented in this report. Samples were prepared in the form of composite sheets made up of small pieces stacked in a pile. Each piece was trimmed with scissors so that an edge was parallel to the rolling direction, dipped in paraffin, and assembled into the pack by aligning it under the cross hair of a microscope. As the desired orientation was obtained on each piece it was secured in place by touching with a hot wire to melt the paraffin. A stack of ten or fifteen pieces was built up in this way, then trimmed to the shape of a T; the portion to be X rayed was then etched to the shape of a wire about 0.045 in. diam with 6N HCl. This method of shaping the sample is a modification of that used by Bakarian on magnesium.&apos; The absorption of the rays in the sample was so slight that it caused no difficulty in interpreting the films. Exposures were made with a 0.030 in. diam pinhole, using molybdenum radiation (40 kv, 25 ma, Type A film at 5 cm, 2 to 3 hr exposures). With the recrystallized specimens it was found necessary to oscillate the specimen so as to reduce the spottiness of the lines. A range of oscillation of 5" was SUB- cient to produce reasonably satisfactory patterns, though the quality was somewhat inferior to that of the deformation texture patterns, and only two degrees of intensity were read from the arcs on the films. Typical photo-grams for each of the deformation textures and the recrystallization texture are assembled in Fig 1. The pole figures were plotted in the usual way with the intensity of the various portions of the diffraction rings estimated by eye. Seven to nine films were made of each sample and each was carefully read in plotting the pole figures. Typical series included exposures with the beam normal to the rolling direction and at 11, 26, 41, 56 and 71" to the cross direction, plus two exposures with the beam normal to the cross direction, and at 11 and 79" respectively to the rolling direction. The rolling was in each case considered sufficient to develop the final texture: the reduction by cold rolling was 84 pct (from 0.0045 to 0.0007 in. thickness), following prior hot rolling in longitudinal and transverse directions and recrystallization; the reduction by hot rolling at 800°C was 90 pct (0.010 to 0.001 in.), following similar prior treatment; the reduction by rolling at 350°C was 88 pct (from 0.005 to 0.0006 in.) after similar prior treatment. The recrystallization texture was determined on a sample rolled at 350" to a reduction of 88 pct (0.0165 to 0.002 in.) after similar prior treatment, then mounted between steel strips to keep it flat and annealed at 700" in an atmosphere of argon. Discussion of Results The results of the X ray determinations are assembled in the pole figures of Fig 2, 3, 4 and 5 for rolling at

Jan 1, 1950

For the third successive year it appears that demand for platinum will exceed supplies of newly-mined metal, forecasts Johnson Matthey in its "Platinum 1987 Interim Review." Based on data available at the end of August 1987, Johnson Matthey estimates that western world demand for platinum could exceed 93t (3 million oz) for the first time ever in 1987. Given a likely increase in demand of 4.3t to around 94t (140,000 oz to around 3.02 million oz), up by about 5% over 1986, there would be a 1.8- to 2.2-t (60,000- to 70,000-oz) shortfall in 1987 supply. South African mines enjoying undisrupted production last year after strikes affected Impala&apos;s output in 1986. The rate of increases in western world supplies, though, has probably not kept pace with the growth in demand. A downturn in US demand for platinum was counterbalanced by substantially increased purchases from Japanese and Western European consumers of the metal. Total imports of platinum into Japan in 1987 may have reached a record level of 45t (1.45 million oz). The largest source of demand for platinum remains its use in autocatalysts US and Japanese auto manufacturers reduced their platinum purchases in the wake of falling car production. Therefore, a small decline in overall 1987 sales of perhaps 780 kt (25,000 oz) was expected in 1987. Western European demand, however, rose strongly as a result of increased output. The spread of exhaust emission control regulations has sharply boosted the number of cars equipped with three-way catalysts containing platinum and rhodium. Western European auto makers probably needed 5.6t (180,000 oz) of platinum last year, a 50% increase over 1986. The jewelry industry, the second largest user of platinum, increased its offtake of platinum in 1987. Japan, the key to the jewelry industry&apos;s consumption of platinum, could absorb 24.8 to 26.4 (800,000 to 850,000 oz) in 1987, compared with 23t (740,000 oz) in 1986. Platinum has wide and diverse applications in industry. Other major users are the chemical, glass, electronics and petroleum industries. They were expected to increase their aggregate demand by about 1.8t (60,000 oz), or 11% in 1987. Investment demand for small platinum bars and coins of up to 311g (10 oz) dropped back in relatively inhospitable conditions. This, after rapid increases in recent years and a surge in US demand in 1986. However, Japanese investors purchasing bars weighing 500g or 1 kg (16 to 32 oz), and considered more speculative, recovered last year. Over the first eight months of 1987, platinum performed strongly on the open market with average prices well up from 1986 in dollar terms. But the increase was less marked where measured in sterling or yen. Platinum ad¬vanced by $3.70/g ($115 per oz) between Jan. 1 and mid-September with the average price having been surpassed only in the heady days of 1980. Johnson Matthey believes that supply and demand appear likely to keep more or less in step on a gently rising path. The recent open market volatility will continue with prices fluctuating widely over short periods. Writing in mid-September, and basing its view on the tangible evidence of platinum&apos;s price fall since the review was written, the company&apos;s assessment of price prospects is unchanged. Demand for platinum remains strong all-round. In October, imports of platinum into Japan were exceptionally high. And in the US and Japan there was a marked resurgence of demand from private investors. Although market sentiment may have been influenced in part by the news of several potential new platinum mines in South Africa, the full effects of these prospects on supply, even if they were all to materialize, would not be felt before the early-mid 1990s, Johnson Matthey said. According to the review, no radical change in platinum supplies is discernible in the near-term. Incremental growth from South Africa can be expected as the major producers continue to make the most of their existing resources. The Soviet Union may be keener to sell than it appeared earlier in 1987, especially if prices hold up, but a substantial increase is improbable. Canadian supplies are not capable of much variation from present levels. The palladium-rich Stillwater mine will make a modest contribution. Referring to the investment market, Johnson Matthey concluded that despite this year&apos;s reduction in sales of small-denomination investment products, platinum does appear to have established itself with private investors. Bearing in mind how business multiplied in comparatively unfavorable conditions before 1986, it is reasonable to look for a resumption of growth once the climate improves.

Jan 2, 1988

By J. D. Anthony

The Australian continent possesses significant reserves of a wide range of minerals, including bauxite, coal, copper, diamonds, gold, iron ore, lead, manganese, mineral sands, nickel, phosphate, silver, tin, uranium, and zinc. Australia&apos;s identified economic resources of many minerals are very large as indicated in Table 1. A sophisticated and highly experienced mineral industry is now an established feature of the Australian economy and Australia is the world&apos;s largest exporter of iron ore, alumina, mineral sands and refined lead and amongst the leading suppliers of many other commodities such as coal, lead and zinc ores/concentrates, nickel, refined zinc, tungsten concentrates and bauxite. The industry exports 70% of its production. This is reflected in the value of Australian mineral exports which have grown from about $200m in 1960/61, comprising 10% of total export receipts, to about $1265m or 29% of export income in 1970/71 to around $7061 representing 37% of Australia&apos;s total export income in 1980/81. Details of the more significant minerals are as follows: Japan (42.1%) USA (11.3%) ASEAN (6.3%) UK (5.9%) F.R. Germany (3.8%) Republic of Korea (3.4%) New Zealand (2.6%) Also see Table 2. AUSTRALIA&apos;S MINERAL RESOURCES POLICIES Federal and State Governments&apos; Responsibilities Australia has a federal system of government comprising six States, a self-governing Territory and a Federal Government. Under the Australian federal system the Constitution sets down the powers of the Federal Government. All powers not assigned to the Federal Government in the Australian Constitution reside automatically with the States. Certain of these broad powers result in the Federal Government having a significant influence on resources development. For example, in being responsible for economic management, the Federal Government&apos;s fiscal and monetary policies have an important effect on industry as well as on State finances. In particular, the taxation regime employed by the Federal Government is of direct importance to decision-makers in the resources industry. The Federal Government is responsible also under the Constitution for external trade matters; and international trade and commodity matters are increasingly important in Australia&apos;s international relationships. Foreign investment is another area where the Federal Government has a role to ensure that national interests are protected. This foreign investment power flows from the Federal Government&apos;s control of foreign exchange movements into and out of Australia. However, before enlarging on these and others of the Federal Government&apos;s powers and policies, it should be emphasized that the State governments, by virtue of their wide powers to regulate matters within their own boundaries, are more directly involved in the day-to-day administration and regulation of mining operations. For instance, the powers of the State governments include the responsibility-for the granting of exploration rights and mining leases, the approval of mining operations and the levying of royalties and other like charges. Administrative arrangements covering the granting of minerals and petroleum exploration and development titles vary from State to State. Before development rights are granted, State governments consider environment protection and rehabilitation aspects of development proposals. The provision of infrastructure within State borders is a matter primarily of State government responsibility. It is usual practice in Australia for State governments to construct and operate infrastructure services such. as railways, ports and electricity generation and transmission. The States may also provide certain public services such as electricity. and water, port and loading facilities, communications, health and education services which form part of the infrastructure of mining operations. In remote areas the mining companies themselves usually are expected to provide much of this infrastructure. However, the Federal Government is primarily responsible in some fields, such as telecommunications and parts of the railways network. State governments carry out preliminary exploration and geological mapping and some are directly involved in the mining of coal for power generation. The Federal Government&apos;s responsibilities in addition to economic management, taxation, international relations, foreign capital and investment, include regulation of exports, environmental matters and matters affecting the Aboriginals of the Northern Territory. FEDERAL GOVERNMENT POLICIES The continued sound development of the minerals and energy resources sector is regarded by the Federal Government as being of very great importance. However, the Government does not seek to participate directly in resource developments. It sees its role rather as that of establishing a sound economic and policy climate in which private companies can identify opportunities, seek out customers and marshall the necessary capital for the development of resource projects.

Jan 1, 1982

By K. E. Nelson

The effect of thorium and zinc variations on the strength and 100-hr creep characteristics of Mg-Th-Zn-Zr alloys was investigated. Optimum resistance to creep at 650° and 700°F are attainable within a certain range of thorium and zinc contents. This range does not conform to that which develops maximum tensile properties. RAPID advances have been made during the last few years in the development of magnesium alloys for elevated temperature applications demanding high resistance to creep. The beneficial effect of rare-earth metals on the creep resistance of magnesium alloys has been emphasized by a number of publications1-13 and such alloys are now in commercial production. The use of thorium as an alloying ingredient in magnesium was mentioned by McDonald and also in two Alien Property Custodian patent applications.10,17 The initial observation of Sauerwald that thorium contributes still higher creep resistance to magnesium than is attainable with rare-earth metals has recently been substantiated.&apos;" " In fact, it has been demonstrated that the useful temperature range of magnesium alloys is appreciably extended by the use of thorium. In all cases, it was observed that zirconium must be included in the alloys in order to render them fine grained and more readily castable. Several recent publications:2-27 indicate that a still further improvement in creep resistance and a further extension of the useful temperature range can be realized by the addition of zinc to alloys. The primary purpose of this paper is to present the results of a comprehensive study of the effect of zinc on the strength and creep characteristics of Mg-Th-Zr alloys. Compositions covering the range of thorium content from M to 6 pct and zinc content from 0 to 5 pct have been investigated. The creep characteristics at 650" and 700°F reported in this paper are based on results of tests of 100-hr duration. It is appreciated that creep tests of 100-hr duration might not yield adequate data for design purposes for parts with much longer expected life. However, for the purposes of the present discussion, it is felt that the combination of stresses and temperatures used in the 100-hr creep tests have yielded a clear representation of the compositional variation of creep resistance at the temperatures investigated. Creep tests of 1000-hr duration are now in progress on a few of the most promising alloys. Preparation and Testing of Alloys The alloys studied in this intensive investigation were prepared in 25 lb capacity mild steel crucibles. The thorium, zinc, and zirconium were alloyed and poured as described in earlier publications The thorium was introduced into the melt in the form of a Mg-Th hardener," the zinc added in the metallic form, and the zirconium alloyed in the form of the commercial hardener containing magnesium and 30 to 50 pct Zr.10, 26 Fluxing practices for melting and refining were the same as for magnesium-rare-earth metal-zirconium alloys. The melts were sampled for analytical determinations and poured into separately cast 1/2 in. diameter standard tensile bars. The test bars were given a precipitation treatment of 16 hr at 600°F in laboratory furnaces. It has been shown by other tests that a high temperature solution treatment followed by an aging treatment is unnecessary for the development of optimum properties in Mg-Th-Zn-Zr alloys. The selection of 600°F as the aging temperature was based on an attempt to achieve metallurgical stability without coalescence of the undissolved phases and the attendant loss in strength. The thorium, zinc, and zirconium contents of each melt were determined chemically. The zirconium contents are reported in two parts, "soluble" and "insoluble," referring, respectively, to the portions present in the alloy which are soluble and insoluble in dilute HCl acid. Distinction is being made between these two components of the zirconium content in the alloys because it has been found that only that portion of the zirconium content which is soluble in dilute inorganic acids affects the structure and properties of the alloys. The usual impurities consisting of copper, iron, manganese, and nickel were determined spectroscopically. The analysis for each melt is listed in Table I. A description of the methods of tension and creep testing has been detailed in earlier papers. The tests were performed with the cast skin re-

Jan 1, 1954

Though the mining industry is beset with problems-a broad slump in metal prices, a rash of acquisitions and takeover attempts, rising capital costs, and an aggressive Canadian energy program-there was an under-current of optimism at the 1981 AMC conference that&apos;s been missing at industry meetings in recent years. The industry anticipates a turnaround in the 1980s, buoyed perhaps by the new spirit of cooperation from the Reagan administration. For the first time in years, it appears that Washington and the mining industry share mutual goals for developing the nation&apos;s resources. A number of prominent Reagan officials were on hand during the three-day conference in Denver, September 28-30, to reassure conferees that restrictions plaguing energy and minerals development would be eased. Interior Secretary James Watt and other government speakers promised cooperation in bringing balance to federal public lands policies, a review of regulations that have hampered the industry, and prompt action on domestic minerals policy. Daniel N. Miller, Jr., assistant interior secretary for energy and minerals, said, "It is our intent to dismantle the overhearing bureaucracy that has stifled mineral and energy exploration. We know that Interior&apos;s rules have shackled the mining industry in recent years, and we know that those rules must be eased." Miller noted that the administration will not permit mining on national park lands or wildlife refuges, but added that "we must assess a variety of uses of public lands ... we will restore lands to multiple use where practical." EPA administrator Anne Gorsuch assured industry representatives that the administration intends to work within environmental laws and with Congress to review and revise water, soil, and air pollution policies. "We have a strong commitment to see that there are major mid-course corrections" of air regulations, Gorsuch said. Support from these "new friends in Washington" will be crucial in light of the serious problems pressuring the industry. Copper prices have plunged to almost three-year lows and prices for molybdenum, gold, and silver have slumped. Earnings are down and today&apos;s high interest rates prevent a quick turnaround. Companies are on edge after the recent takeovers of Kennecott, St. Joe Minerals, and Texasgulf Though the companies may benefit from the major injection of capital that large energy firms can provide, many observers fear the new parent companies won&apos;t be willing or able to survive the cyclical metal markets. Arco and Gulf Resources, for example, have been criticized for closing Anaconda&apos;s Montana copper facility and the Bunker Hill unit in Idaho, respectively. Canada&apos;s national energy program has brought additional pressure on the industry. This policy, which emphasizes Canadian majority ownership of the country&apos;s energy sources, resulted in acquisition of Texasgulf&apos;s Kidd Creek lead-zinc-silver complex by Canadian and French interests. In response, the American Mining Congress adopted two resolutions on takeovers. One calls for the US government to review laws regulating tender offers and takeovers "to better ensure that principles of fairness and sound investment policy govern such transactions.&apos;&apos; The other asks the government to demand "fairness and reciprocity" as a basis for US policy governing international trade in minerals and investment relations with other countries. AMC&apos;s shopping list included four other priority goals, aimed primarily at the Interior Department: • Prompt implementation of Interior&apos;s efforts to bring order and balance to rules governing access to public lands for exploration and development. • Statement by the President of a comprehensive and coherent domestic minerals policy. • Amendment of the Clean Air Act to remove impediments to productivity and barriers to economic development that are not necessary to essential goals of the act. • Development of new approaches to better achieve worker health and safety in the most cost-effective manner without penalizing productivity. Almost 6000 mining industry executives attended the 1981 convention-the largest in AMC&apos;s 84-year history. As in years past, the conference offered a wide spectrum of presentations-nearly 100 papers in 26 sessions-ranging from commodity forecasts to the latest technical achievements. New Administration Stockpile Goals The minerals availability session included a presentation on management of the minerals stockpile by Paul K. Kruege r of the Federal Emergency Management Agency. Krueger notes that rather than providing incentives for development of a domestic raw materials industry, the government has "a history of providing disincentives to industry." In mining industries, he adds, the rate of productivity growth has been -6% over the last six or seven years. Krueger feels there will be a major change under the Reagan administration. In March 1981, FEMA began the first purchase program for the national defense stockpile in over 20 years. The stockpile is sadly out of balance, Krueger adds. Stockpile materials are

Jan 12, 1981

By R. E. Hudrlik, G. W. Preckshot

The diffusion coefficients of silver from solid silver in molten lead were measured to within ± 0.8 pet in a columnar type diffusion cell ower, the temperature range of 326° to 530°C. Fick&apos;s law describes the process up to 530°C where the laminar mechanism appareltly breaks down. These is negligible resistance at the interface as shown by mathematical analyses. The diffusion coefficients are found concentration independent. IT would seem that diffusion in liquid metals would be free of such effects as molecular structure, dissociation. polarization. and compound formation. This view was taken by Gorman and preckshot in their study of diffusion of copper from solid copper into molten lead. They reported diffusion coefficients which were independent of the concentration over the range of 478° to 750°C. They found that the Stokes-Einstein equation with constant radius of the diffusing specie represented the diffusion data better than Eyring&apos;s rate theory equation and Sheibel&apos;s correlation. The radius of diffusion was found to be that of the doubly charged copper. There appeared to be no resistance across the solid-liquid boundary. In the present work the diffusion coefficients for silver in liquid lead were measured over a range of temperatures of 350° to 505°C. The solubility of silver in lead over the range of 303° to 630°C was also obtained. These results are compared with calculated or correlated values or with data in the literature. EXPERIMENTAL Procedure—The experimental equipment techniques and procedures were those reported in detail by Gorman and preckshot9 and will not be repeated here. Measured values of WT, Co, A. L were obtained for various diffusion times and the diffusion coefficient was computed for the case of no resistance at the interface9, 11 by: WT/CoAL = 1- 8/p2 n=1 1/(2n - 1)2 exp[-(2n - 1)2p2 Dt/4L2] [1] or where there was resistance at the interface by: WT = 1- ?n=1 2h2/ap2L [sxp [-Dan2t]/[(h2 + an2) L + h] The roots an are those of the transcendental equation3 tan (an L) = Iz/cun. The diffusion coefficient is that defined by Hartley and Crank.7 The total silver in the lead cylinder and equilibrium slug was determined by a cupellation technique&apos; with proper correction for losses. Analysis of known samples showed that this method is surprisingly accurate. The amount of silver in the lead adhering to the silver cylinder was obtained in the same fashion as shown by Gorman and preckshot.9 The small errors involved in this determination are not critical since the silver in this adhering lead layer is only 3 to 15 pet of the total diffused. Materials—Electrolytic silver containing 99.9+ pet Ag obtained from General Refineries of Minneapolis, Minn. was used for all but runs 7 and 8. For the balance of the runs this silver was reduced with hydrogen at 1100°C and its oxygen content was found to be about 0.017 pet. For the runs. 7 and 8, phosphorous-reduced silver of the same purity was obtained from Handy and Harman Co. of Chicago, Ill. The densities of the phosphorus-reduced silver and the hydrogen-reduced electrolytic silver were 10.484 and 10.487 g per cm3, respectively. These values agree with those reported for pure silver. Silver which was reduced at 900 C had an average density of 9.998 g per cm3, indicating porosity. This silver was used for a number of runs which were not tabulated in Table I. These are indicated by crosses on Fig. 2. The 99.999 pet Pb was obtained from the National Lead Co. Research Laboratory of Brooklyn, New York. DISCUSSION OF RESULTS The diffusion and solubility results are reported in Table I for eleven runs using either phosphorus-reduced electrolytic silver or hydrogen-reduced silver at 1100° C. The solubility data shown in Fig. 1 show the excellent agreement with that reported by Heycock and Neville.8 The data of Friedrichs5 apparently are in error. The experimental solubility data of this work are reported to 0.3 pet. The experimental diffusion coefficients computed from Eq. [1] are reported within 1.2 pet of the mean and are plotted in Fig. 2. These are expressed within +0.8 pet of the experimental values over the entire temperature range by: D= 8.26 x 10 -5 e-1925/RT . [3] There appears to be little difference due to the

Jan 1, 1961

By T. Watanabe, S. Karashima, H. Oikawa

Creep tests of a! iron and its solid-solution Fe-Mo, Fe-Co, and Fe-Si alloys with bcc structure were conducted under constant stresses in ferromagnetic and paramagnetic temperature regions above 0.5T,(T, is the absolute melting temperature). It was found their high-temperature creep behavior changed in the vicinity of the magnetic transformation temperature, that is Curie temperature, TC. In the ferromagnetic region below the steady-state creep rates were lower than those expected from the extrapolation of the data in the paramagnetic region, the amount of decrease being affected by solute addition. Also, activation energies for steady-state creep were different in the two temperature regions. The observed changes were concluded to be due to the effect of ferromagnetism on diffusion. It is well-known that plastic deformation of metals and alloys in the temperature region where mobility of atoms becomes high enough is controlled by atomic diffusion." Many experimental results394 on fcc and hcp materials have been reported to confirm this view. In bcc metals and alloys, not much basic investigation of creep deformation has been performed. In particular, experimental work on iron5-&apos; and its alloys, covering a quite wide temperature range is extremely scarce. On the other hand, the effect of ferromagnetism on diffusion has recently been demonstrated in a iron and its alloys;&apos; it has been made clear that diffusion rates in ferromagnetic temperatur~e region are significantly lower than those expected from diffusion measurement in the paramagnetic region. Therefore, it is concluded that the magnetic effect influences the high-temperature creep behavior in a! iron and in its alloys. The present authors have conducted creep experiments on a iron and its solid-solution alloys over a very wide temperature range, and have demonstrated the magnetic effect. A part of the result has already been reported elsewhere.26 In this paper, the experimental results will be described in detail. While the manuscript was in preparation, a similar magnetic effect on creep deformation of a iron was reported by Ishida . Though their experiments covered an extremely wide temperature range, the creep stresses varied from several thousand psi to several ten thousand psi with decreasing temperature. Consequently, it may be said that there remain some questions concerning their results. I) EXPERIMENTAL PROCEDURE The materials used in the experiments were pure iron and its alloys; they were prepared by vacuum melting (10"" mm Hg) electrolytic iron (99.9 pct), molybdenum powder (99.9 pct), ultrapure silicon, and cobalt pellet (99.5 pct) in alumina crucibles. The ingots were hot-forged to plates 10 mm thick and 50 mm wide, and then were hot-rolled at about 700°C into sheets 1 mm thick. Creep specimens, 5 mm in width and 15 mm in gage length, were machined from the sheets. They were annealed at temperatures which were well above their respective recrystallization temperatures. The chemical compositions of the metal and alloys are shown in Table I together with their heat treatments. High-temperature creep tests by the use of a lever-type creep testing machine were carried out in argon atmosphere under constant tensile stress. In order to keep &eep stress constant within about *0.5 pct, stress change due to specimen elongation was compensated by tensile force of a stainless-steel bellows which was inserted between the loading lever and lower specimen grip with the assumptions of a linear relationship between change in cross section and strain and uniform strain along the section. Beyond the limit of compensation by the bellows, small amounts of load were removed after appropriate strains. The testing temperatures were maintained to within i2"C of the reported values. Creep deformation was auto-graphically recorded at the upper (moving) specimen grip using a linear differential transformer which was held by a stainless-steel rod mechanically connected to the lower specimen grip. It was also directly measured with a dial-gage reading to -& mm. 11) EXPERIMENTAL RESULTS AND DISCUSSION 1) Creep Curves of a Iron and Its Alloys. Creep curves of a! iron, Fe-Si, Fe-Co, and Fe-Mo (<2 at. pct Mo) alloys usually consisted of three stages, that is transient (primary), steady-state (secondary), and accelerating (tertiary) stages. An example is illustrated in Fig. 1 for a! iron. However, in Fe-Mo alloys with more than 2 at. pct Mo, an inverse transient creepZ8 was observed as indicated in Fig. 2. The inverse behavior, which varied with the amount of molybdenum, may be due to the substructure* in the annealed specimen as was sug-

Jan 1, 1969

By K. C. Liddell

Chemical processing research made notable advances during 1986. This overview is necessarily brief. It is intended, though, to give the flavor of selected areas where recent activity has been high. At the 1986 Annual Meeting of The Metallurgical Society, a symposium was held on hydrometallurgical reactor design and kinetics. R. G. Bautista, R. J. Wesely, and G. W. Warren edited the resultant proceedings volume. This is the first available volume on this subject. It includes 26 papers on fundamental kinetic studies, modeling of reaction kinetics and reactor performance, and pilot and production-scale operations. Metals discussed include silver, gold, platinum, titanium, cobalt, nickel, zinc, copper, manganese, and iron. Agitated tanks, heaps, dumps, pachucas, pressure auto-claves, electrolytic reactors, liquid membranes, and fluidized beds are among the reactor types covered. An international symposium dealt with iron control in hydrometallurgy. It was sponsored by the Metallurgical Society of CIM, the Institution of Mining and Metallurgy, The Metallurgical Society of AIME, and Gesellschaft Deutscher Metallhutten and Bergleute. J. E. Dutrizac and A. J. Monhemius edited a proceedings volume containing 41 papers. These covered many areas of process chemistry, solvent extraction, precipitation, treatment of pickle liquors, impurities, residues and the environment, and process selection. Precious metals Concerning precious metals processing, the US Bureau of Mines IC 9059 is noteworthy. It includes papers presented at a USBM briefing at the Western Mining Conference in Denver. Topics covered included ion exchange, staged heap-leaching direct-electrowinning, and mercury precipitation during cyanide leaching of gold ores. Also dealt with were carbonaceous gold ores, carbon adsorption and desorption, heap leaching, the carbon-in-pulp process, and precious metals recovery from electronic scrap and solder. R. C. Sandberg and J. L. Huiatt (Journal of Metals, June 1986, and USBM RI 9022) developed a method to recover silver, gold, and lead from a complex sulfide using ferric chloride, thiourea, and brine-leach solutions. Gallium Gallium recovery was also a subject of considerable interest. Much information on this topic is proprietary. But two papers described solvent extraction of gallium. V. P. Judin and R. G. Bautista (Metallurgical Transactions B, 1986) developed an equilibrium model to separate gallium chloride from aluminum chloride. Tributyl phosphate was the extractant studied. T. Sato and H. Oishi (Hydrometallurgy, 16, 1986) investigated gallium extraction from sodium hydroxide by using Kelex 100. Data were given on the equilibrium distribution and the extraction kinetics. Galena Interest continues in recovering lead by hydrometallurgical processing of galena. J. E. Dutrizac; S. H. Kim, H. Henein, and G. W. Warren; and M. C. Fuerstenau et al. (Metallurgical Transactions B, 1986) all investigated leaching of PbS by ferric chloride. Dutrizac reported parabolic kinetics and reaction control by outward diffusion of lead through a porous layer of elemental sulfur. Kim, Henein, and Warren, however, reported that two leaching reactions occur. A nonoxidative reaction produces H2S and there is an oxidative leaching by ferric ion. Acid dissolution was found to predominate when the acid activity to ferric activity ratio is high. Surface chemical reaction was reported to be controlling. Fuerstenau et al. also reported that mass transfer through the sulfur layer is rate determining. But they believe that the species transferred are chloro complexes of ferric ion. In other work on lead, A. Y. Lee, A. M. Wethington, and E. R. Cole, Jr. of the USBM described an environmentally acceptable hydrometallurgical alternative to the smelting of lead concentrates (RI 9055). Uranium In spite of depressed prices, much work has been reported on uranium processing. R. G. L. McCready, D. Wadden, and A. March-bank (Hydrometallurgy 17, 1986) described the nutrient needs for in-place leaching by T. ferrooxidans. They reported the optimal conditions for uranium solubilization. L. E. Eary, H. L. Barnes, and L. M. Cathles (Metallurgical Transactions B, 1986) carried out an experimental and modeling study of uraninite dissolution. They concluded that ferric ion preferentially leaches uraninite in pyritic ores. Hydrogen peroxide was found to be less selective. A. Vuorinen, P. Hiltunen, and O. H. Tuovinen (Hydrometallurgy 15, 1986) studied redox and precipitation reactions of iron and uranium in leach liquors. P. T. Chiang (Hydrometallurgy 17, 1986) reported on the effect of uranium loading in the DEPA-TOPO process to separate uranium and iron from wet-process phosphoric acid. F. J. Hurst (Hydrometallurgy 16, 1986) conducted a fundamental study of the separation of uranium from phosphoric acid by DEPA and TOPO. D. E. Chia and W. C. Cooper (Hydrometallurgy 16, 1986) reported on bench and pilot scale work to recover uranium by the HIMIX process. They reported that the acid consumption to produce an eluate suitable for yellowcake

Jan 5, 1987

By G. A. Swanquist, E. M. Morales

INTRODUCTION The idea of water management and control at the Church Rock Mill operations began to take shape in February 1979. At that time, we were already investigating the feasibility of decreasing the fresh water requirements so that the solids would become the limiting factor in tailings impoundment utilization. The area for solution evaporation could be kept at a fraction of the normal requirements under the standard process of full water usage. The Church Rock Mill is an acid leach circuit followed by solids/liquid separation with thickeners in counter current decantation, and solvent extraction. Following the normal design of acid leach circuits, reuse of tailings solution was not incorporated in the original mill process design. INITIAL WATER CONTROL INVESTIGATIONS The investigations to decrease the fresh water requirements centered around modifying the grinding circuit from the present semi-autogenous grinding (SAG) mill in closed circuit with hydrocyclones, to open circuit grinding with a rod mill. The open circuit grinding with the SAG mill and rod mill in series had the potential of decreasing the water requirements for grinding and leach dilution by approximately 50% or 1.4 m3/min (300 gpm). The grinding pulp density would be maintained at 70 to 72% solids, and the leach dilution to 50% solids would be accomplished with acid tailings liquor recycle. In such a grinding circuit arrangement, the SAG mill would provide the primary or coarse grind, and the rod mill would be used for the fine grind. By the SAG mill and rod mill series grinding method of water control and other secondary water controls in various places downstream from the grinding circuit, the required necessary evaporation area was estimated at 120 acres of liquid surface. A second method of water control at grinding was investigated. A two-stage cyclone classification circuit appeared to have a good potential of achieving the same water reduction at a much lower capital and operating cost. However, in retrospect, this would not have been a viable method since a high slime recycle load would have been established hindering classification. The use of reagents to neutralize the acid tailings solution was not considered seriously at that time, since it would have materially increased operating costs, although it would have also allowed more tailings solution recycle and consequently, less fresh water usage. However, with the tailings solution deposition area available at that time, it was not then necessary to incur the high cost of neutralization. The control expected by the series grinding of semiautogenous and rod mills would have been sufficient to maintain a water consumption/evaporation equilibrium well in line with the available land area. IMPLEMENTATION OF NEUTRALIZATION OPERATIONS During the summer of 1979, the UNC Church Rock Mill experienced a tailings dam breach which resulted in a prolonged mill shutdown. Upon resumption of operations at the end of October 1979, tailings deposition was restricted to a small portion of the tailings impoundment area. Figure 1 shows the general tailings area and the limits of the present deposition area in the central part including the borrow pits. These borrow pits had been excavated to provide materials for tailings dam construction. Immediately after resumption of operations, it became evident that it would be necessary to control the quantity of liquid to be evaporated because of the small confined area available for tailings solution deposition and to maximize the deposition time in the tailings area. The water control required had to be exercised on a large scale, and to be in operation as quickly as possible. An obvious solution was to reuse the tailings liquor in mill process. Immediate steps were taken to install the necessary equipment for tailings neutralization on an interim basis. Anhydrous ammonia was selected as the primary neutralization reagent since it was the quickest system that could be placed in operation. Previous laboratory tests indicated fair results with ammonia neutralization. Such a system required a minimum of installed equipment and handling. INITIAL NEUTRALIZATION OPERATIONS Actual neutralization operations began on November 26, 1979. The raffinate solution which normally would have been discarded was pumped to a 3.7 m (12ft) diam by 4.3 m (14ft) tank for reagent contact, see Figure 2. At this tank, anhydrous ammonia was added directly from the tanker trailers and controlled at pH 7.0 nominally. Agitation was provided by air sparging. The neutralized product formed a highly viscous slurry in the grinding circuit which resulted in pumping and cyclone classification problems.

Jan 1, 1982

By M. Bevis, A. F. Acton, P. C. Rowlands

Defor~nation twinning and transformation twinning modes most likely to be operative in Fe-Ni and Fe-Ni-C martensites have been determined using a new theory of the crystallography of deformation t~inning.~ This analysis shows that potentially important conventional and nonconventional twinning modes1 have been omitted in previous analyses. Discussion is given on the relevance of the predicted twinning modes to the lattice invariant shear associated with the martensite transformation in steels and to anomalous deformation twinning in Fe-Ni-C martensites. THE two most important criteria which appear to govern operative twinning modes in metallic structures1 are that the magnitude of the twinning shear should be small and that the twinning shear should restore the lattice or a multiple lattice in a twin orientation. The latter criterion ensures that the shuffle mechanism required to restore the structure in a twin orientation is simple. These criteria have been adhered to in the prediction of twinning modes2"6 in bcc and bct single-lattice structures with axial ratios in the range y = 1 to 1.09 as, for example, encountered in martensite occurring in steels. Refs. 2 and 3 in particular consider the martensite transformation in steels and the twinning modes in these cases relate to transformation twinning, and hence the lattice invariant shear associated with the martensite transformation. The list of twinning modes which can be compiled from these sources is incomplete and the ranges of magnitude of shear considered could be unrealistically small, particularly in the case of deformation twinning. The latter consideration is supported by the fact that twinning modes with magnitudes of shear large compared with the smallest shear consistent with a simple shuffle mechanism have been established in, for example, the single-lattice structure mercury7 and the multiple-lattice structure zirconium.&apos; In addition the anomalous deformation twins reported by Ftichrnan4 to occur in a range of Fe-Ni-C martensites still remain unexplained. It is clear that a comprehensive analysis of twinning modes likely to be operative in martensite In steels is required. The results of the application of a new theory of the crystallography of deformation twinningg to these structures are presented in this paper. The theory has been used to determine all shears which restore the lattice or a multiple lattice in a new orientation with magnitude of shear up to a required maximum. The orientation relationships between parent and twinned lattices are not restricted to the classical orientation relationships of reflection in the twin plane or a rotation of 180 deg about the shear direction. PREDICTED TWINNING MODES Twinning modes which restore all or one half of lattice points to their correct twin positions will be referred to as m = 1 and m = 2 modes, respectively. These modes are the most likely to describe operative modes in single lattice structures. The bcc m = 1 and m = 2 modes which have magnitudes of shear s in the range s < 2 and s < 1, respectively, have been given10 and are reproduced here in Tables I and 11. Detailed discussion of the crystallography of these modes and cubic modes in general will be discussed elsewhere (~evis and rocker, to be published). The four twinning elements Kl, &,ql,7)2 as well as the magnitude of shear s are given for each twinning mode, and the twinning modes are given in order of increasing shear. Two twinning modes are given in each row of the tables, the twinning mode Kl, Kz, ql, q2 and the reciprocal twinning mode with elements Kl = K,, Ki = Kl, q: = q2, and 17; = ql. The m = 1 and m = 2 twinning modes which describe twinning shears with small magnitudes of shear and simple shuffle mechanisms in bct crystals with -y = 1 to 1.09 are given in Tables I11 and IV, respectively. On increasing the symmetry of the tetragonal lattice to cubic, that is making y = 1, all modes listed in Tables 111 and IV must reduce to crystallographically equivalent variants of the modes given in Tables I and 11, respectively, or become twinning modes with both shear planes as symmetry planes in the cubic lattice and hence not considered in Tables I and 11. With the exception of this last type of mode only those tetragonal twinning modes which reduce to modes 1.1, 1.2, 2.1, and 2.2 of Tables I and I1 are considered in Tables 111 and IV. For values of y in the range -y = 1 to 1.09 the tetragonal modes and the corresponding cubic twinning modes have approximately the same magnitude of shear. The twinning modes listed in Tables 111 and IV are therefore by the criteria given above the most

Jan 1, 1969