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By Fumihiko Sumiya, Kunihisa Katsuyama, Yuji Ogata, Yuji Wada, Yukio Katoh, Yoshikazu Hirosaki

In our previous work, the results of photographic observation using a high speed framing camera showed that the difference between precursor air shock wave velocity and detonation velocity was primary factor for the channel effect. It is assumed that the decreasing precursor air shock wave velocity can prevent the channel effect in charge hole.

Jan 1, 1996

The economics of aerial survey and the technical processes by the aid of which vertical and oblique aerial photographs are turned into line maps showing the most profuse topographical detail such as contours, rivers, roads, plantations, habitations and cultivated lands do not concern us here. They have been ably and exhaustively described in the publications of the British War Office, handbooks on surveying methods and on photogrammetry and in the South African mining, engineering and geological press. Suffice it to mention that for geological interpretation vertical photographs are used on such scales as 1/5000 to 1/10,000, overlapping each other by 55 to 60 per cent in the direction of flight and some 10 to 20 per cent sideways on to the next sequence or strip. Through this duplication in the overlap the stereoscopic examination becomes possible. The standards by which monochrome aerial photographs are interpreted can be grouped under the following headings: 1. Continuity.—The continuity of topographic relief not obtainable on the ground owing to the smallness of man relative to that relief. Here lies the greatest value of aerial survey with relation to geology. An immediate alignment can often be made of drainages running in different directions on each side of a hill; saddles and gorges in the mountainsides; abrupt termination of outcrop or indubitable signs of faulting. Most conclusive evidence is thus often obtained about faults which may be perfectly visible at one place on the ground but cannot elsewhere be identified because of decreased throw. Similarly, dikes can often be followed notwithstanding frequent interruption and disappearance under the soil. Scattered outcrop in scrub-covered or bush-covered country can be conveniently located and often correlated, with a large saving of time compared with field work. The same applies to geological contacts. Once these have been established at a few points on the ground, the intervening stretches can be drawn in.

Jan 1, 1937

By A. C. Atenas

The majority of refineries that operate with traditional starting sheet technology ?starting sheet- uses a relatively low current density of 250 A/m2. Industrial tests at high current density with conventional technology were carried out in Refinery of Potrerillos, getting satisfactory results that are similar the results obtained with permanent cathode technology. The series of copper refining electrolyses electrolysis trials were carried out between 300 to 325 A/m2. Electrolysis conditions, in accordance with industrial practice, were kept constant in all sets of experiments and were as follows; 45 g/L Cu 2+, 210 g/L H2SO4, 62 °C, electrolyte circulation of 14-18 L/min. The same Dosing dosing of Additives additives used in permanent cathode technology was required in these tests. In general, only not importantminor operational modifications were necessary in comparison to normal practice in the Refinery to assure the good results. Different quality of starting sheet and anodes were examined. The test included starting sheets from smelter and laminated process. Results indicate that the charge physical quality of the electrodes and the control of the process during the two first days of operation are the keys to get the best results.

Jan 1, 2007

It is desired to call the attention of young men who have had technical training and experience to the fact that their abilities can best be put at the service of the country by selecting a branch of service in which their special qualifications will be of the greatest use. The submarine force of the United States Navy requires the services, as officers on board submarines, of young men who have had technical training in mechanical and electrical engineering and who have had experience in these professions. It is intended to enroll a number of such men as provisional ensigns in the Naval Reserve Force, give then a course of instruction in deck duties at Annapolis and a course in submarine work at New London. Those who successfully pass these courses will then be sent on hoard submarines for regular duty. It is requested that any men who desire this duty and who are qualified as below outlined send their names and addresses to the Commander Submarine Force, U. S. S. Chicago, care of Postmaster, New York. Qualifications required: Desire to serve in submarines. Degree of M. E. E. E., or E. M. Two and one-half years' practical experience in profession. Not over 35 years old. Physically strong and sound.

Jan 7, 1918

By A. J. Burger, F. L. D. Cloete

Tests indicated that effective recovery of pyrite from a coarse Witwatersrand quartzitic ore can be accomplished by intensified flotation in an air-sparged hydrocyclone. Air is fed through the porous cylindrical wall of the hydrocyclone into a swirling layer of slury. Hydrophobic particles are attached selectively to small air bubbles, which form a froth and are recovered through the overflow. Hydrophilic particles pass out in the underflow. A spigot in which flow is restricted by an orifice was found to be more successful in preventing blockage by wood chips than an annular opening between a pedestal and the cylinder wall. Pyrite recoveries of 85 to 93 per cent at grades of 35 to 40 per cent sulphur were obtained in the air-sparged hydroclone with residence times of about 1 second. Comparative tests in a batch flotation cell after 320 seconds yielded average recoveries of up to 96 per cent, but with a mean grade of 32,5 per cent sulphur. A high colector dosage of 120 g/t was required in the cyclone, compared with the 40 gft used in conventional operations. An air flowrate of about 150 I/min was adequate at a slury feed rate of 35l/min. Particle sizes between 38 and 106µm yielded the best flotation results.

Jan 1, 1988

By Thomas J. O’Neil

From the time of its emergence as an independent state until the late 1950's, the Republic of Ireland suffered from chronic unemployment, the lowest living standard in Northern Europe, and-most serious of all-an aggravated emigration problem which drained off most of its talented young people. In 1958, however, there was a drastic legislative reaction to these problems, and this in turn ignited the current mining boom in Ireland. Paving the Way to Prosperity In order to reverse the tide of emigration, the Irish government now realized, challenging domestic employment opportunities had to be increased and living standards had to be raised. It was also apparent that future industrial and economic growth in Ireland had to be based on producing goods for export. The home market could not sustain large-scale expansion, and consequently Ireland's chronic balance-of-payments deficit could best be minimized by producing export goods. These considerations, together with the limited amount of accumulated wealth in Ireland, combined to set the objectives of the country's First Program for Economic Expansion in 1958. The general objectives of this program were (1) to stimulate all segments of the economy, (2) to develop foreign trade and (3) to attract foreign capital.

Jan 1, 1970

By Ritchie I

Copper minerals are often associated with gold ores. These create processing problems since cyanide solutions are not selective for gold alone. Copper minerals are also leached by cyanide, and typically each one per cent of `reactive' copper in the ore consumes about 30 kg t 1 cyanide. Once in solution, copper cyanide species create further processing problems as they adsorb onto carbon along with the dicyanoaurate ion. These problems make conventional cyanidation of gold ore uneconomic if the ore contains too much copper. As a result, methods of economically processing gold-copper ores are becoming of increasing importance.One of the more promising methods of treating these ores is by selective leaching of the gold by an ammoniacal cyanide solution, a method first described by Hunt in 1901. This method is already in commercial use in Mauritania, treating copper/gold tailings. In general terms, the addition of I - 2 g U1 ammonia to a standard cyanide leach solution causes a significant increase in gold extraction and decrease in copper extraction. However, early testwork on this system has concluded that the optimal leach conditions are critically dependent on ore composition, and the mechanism is not well understood, making predictions of best conditions impossible.

Jan 1, 1997

By J. F. Myers, F. M. Lewis

DISCUSSION L. E. Djingheuzian (Canadian Dept. of Mines and Technical Surveys, Ottawa)—In their Summary the authors say: "Reconciling the grinding efficiency with good metallurgy is still a problem." In the discussion of the first paper8 in his reply to W. I. Garms, Mr. Myers states: "Our grinding process with smooth I-in. balls has reduced by nearly one half the metallic losses in the fine micron sizes of the tailing. This is simply because less of the fine micron sizes are produced. Since the + 65 mesh size is the same as formerly, a higher percentage of the intermediate sizes are developed. These sizes have the highest floatability, require the least reagents, and use less floating time. "These factors contribute so heavily to the overall economies that dropping our power grinding gain from 28 pct back to 19 pct is a small detail. However, we feel that this is only a momentary situation and that eventually the best features of the grinding and flotation processes can be brought together, which is as it should be." Italics are mine. The above statements, to me, appear to be the answer to the opening statement in the Summary. Denoting the costs at different power grinding gains as: Power Grinding Power Grinding Gain, 28 Pct Gala, 19 Pct Cost of grinding G G1 Cost of flotation F F1 Value of metallic losses T T1 where G1 > G2 F3 < F, and T1 < T, we have: G1+Fl+T1<G +F+T. Since the authors accept the idea that "grinding in flotation plants becomes part of the &apos;conditioning&apos; of the feed to flotation",4 i.e., that in flotation the ball mill is primarily a conditioning machine, it can be postulated that Tennessee Copper grinding at cost G1 is more efficient than grinding at lower cost G. This can be directly inferred from the Conclusion of the paper. Mr. Myers also emphasizes this at the end of his reply to Mr. Garms: "that grinding is for the purpose of preparing flotation feed and not grinding per se." This, to me, in the final analysis means that when the efficiency of grinding is weighted against the conditioning factor, the former becomes a function of efficient conditioning, hence, within the system in which proper conditioning is the dominant factor, the best grinding efficiency is provided by grinding which will contribute towards the optimum conditioning. This brings us again to the statement: "that if every grinding unit were considered as a conditioner for each following step, efficient grinding plants would become much easier to design."&apos; In other words, grinding equipment should be balanced against the flotation equipment and against chemical reactions taking place in the system. F. C. Bond (Allis-Chalmers Mfg. Co., Milwaukee)— The authors&apos; discussion of the probable ball motion in a slow speed high dilution mill is very interesting. When the 1-in. balls have worn down to about one fourth of their original weight they apparently first develop a flat surface; as wear progresses this flat face becomes concave, and other concave faces appear. It seems more probable that the first flat face may form at the softest part of the ball surface, and that each succeeding contact tends to force this flat face into sliding contact with a larger round ball; than that the flat faced ball tends to pair off with a particular round ball and to travel with it continuously. When the small worn ball has a flat face and is in sliding contact with a large round ball, the surrounding large balls will assume a more or less definite pattern, and slide against the worn ball, thus producing secondary concave faces. The primary concave face seems to be larger and better developed than the secondary faces. The ball charge can be divided into "concaves" which show at least one concave surface, "intermediates" which have developed flats or incipient concaves, and "rounds." Ball slippage is always present in a tumbling mill, and the mutual ball movement is necessarily a combination of rolling and sliding. The sliding motion is apparently concentrated upon the smaller worn balls which nest between the surrounding larger round balls. When each worn ball starts its upward path in the mill its primary flat or concave surface fits against a larger round ball, and the round ball slides upon it. The action may be something like that of the ball separator in a ball bearing, except that the worn sliding balls are always under considerable pressure. The material is ground under the combined influence of breakage 1—by impacts between falling balls and between falling and supported balls, 2—by being nipped between rolling balls, and 3—by being rubbed between the sliding balls. The rubbing action will be increased in the presence of worn balls with concave surfaces. The rubbing action probably produces a considerable portion of the finely ground slimes in the product. The worn balls commonly approach tetrahedrons in shape, and are very different from concavex, each of which has two equal opposed concave surfaces. Concavex were designed only to grind upon themselves, and not for use in combination with grinding balls. Their action in a grinding charge is very different from

Jan 1, 1952

By R. J. Kuczynski

The U.S. Bureau of Mines Reno Research Center investigated, developed, and patented a high-temperature cyanide leaching process for recovering platinum-group metals (PGM) from automobile catalysts. A batch pilot plant was constructed at the center and operated to demonstrate this technology to industry. Approximately 1,600 kg of used pellet, monolith, and metal support catalysts containing 1,700 g (54 tr oz) of PGM was processed. Forty-five-kilogram batches of used catalysts were leached with sodium cyanide solution at 160°C for 1 h in a countercurrent processing sequence. This selectively dissolved the PGM. Average extractions from the used pellet catalyst were 96 pet Pt, 95 pct Pd, and 73 pct Rh. Average extractions from the used monolith catalyst were 84 pct Pt, 81 pct Pd, and 66 pet Rh. Heating the pregnant leach solutions to 275°C for 4 h destroyed the PGM-cyanide complexes, causing over 99.7 pct of the PGM to precipitate from solution. The PGM precipitate was predominantly a metallic powder concentrate, typically analyzing greater than 50 pct PGM. Heating destroyed the cyanide to less than 0.2 mg/L free and total cyanide. Both the pellet and monolith residues were evaluated for disposal using the toxicity characteristic leaching procedure developed by the U.S. Environmental Protection Agency.

Jan 1, 2010

By A. Ohno

"A new near net shape casting process, OHNO Continuous Casting (OCC), has been recently developed. In this process, molten metal is fed into a heated mold maintained at a temperature above the solidification temperature of the metal being cast. Solidification begins from the central core of the casting rather than the wall of the mold. With the heated mold, the formation of crystals on the mold wall is prevented and solidification occurs out of contact with the mold surface. This produces near net shaped products with smooth, mirror-like surfaces and unidirectional grain structures free from cavities, porosity and segregation. Any crystals formed at the beginning of solidification decrease in number as solidification progresses and eventually a single crystal product is obtained. This process is applicable to any metal alloys, converting them into new materials with superior properties and quality.IntroductionAs a result of growing technical demands for more advanced and diversified materials, the OHNO Continuous Casting (OCC) technology has found many new applications in industry. Considerable interest in the process has been shown in Japan and other countries because of its distinct ability to:produce a single crystal, (1)produce a remarkably clean cast surface without a witness mark,provide extremely good ductility,and reduce casting defects and segregation.This paper provides an overview of the principle of the OCC process and describes the remarkable characteristics of the material produced by this method."

Jan 1, 1988

By Dirk Verhulst, V. I. Lakshmanan

Why Chloride Hydrometallurgy?The most available, cheapest and best known reagent for treating ores and other raw materials for the metallurgical industry is sulfuric acid. Hydrochloric acid is second in line. What often happens is that a process is developed on the basis of sulfuric acid, and that at some point the question arises: wouldn't hydrochloric acid do this better?Considering hydrochloric acid immediately opens a very interesting field of research. To summarize a few comments from another paper presented at this symposium [1]: chlorides are far more soluble than sulfates (with a few interesting exceptions); the hydrogen ion activity is greater in chloride than in comparable sulfate solutions; higher oxidative conditions are possible; regeneration is more straightforward; more diversity is possible and better conditions for the separation of species can be found for solvent extraction and ion exchange; sulfur in the ore can be separated as elemental S; higher metal concentrations in solution and more flexibility in the chosen oxidation state means significantly lower energy costs.Also, the volatility of many chlorides at higher temperature allows further separation without the need to go to a different system. Chlorides also form molten salts with a low melting point, providing another means of separation and purification in the same chloride medium.

Jan 1, 2011

By Orville R. Lyons

DISCUSSION Judson S. Hubbard (The Humphreys Investment Co., Denver)—In this very interesting paper several brief references are made to the Humphreys spiral, a device used for cleaning fine coal. In Table I, Plants 76 and 77, data are given on spiral performance treating Raton and Trinidad coal. The fine coal, as fed to the spiral in these instances, is actually a table middling, hence the more easily treated material was previously removed and a large number of particles were present which were difficult to clean. Mr. W. M. Bertholf of the Colorado Fuel and Iron Co. presented a paper in February, 1946, Cleaning Table Middlings from a Coal Washery with the Humphreys Spiral Concentrator, from which I quote: "In considering the results of our tests it should be noted that the feed was our table middling, and that any real separation is a &apos;moral victory&apos; as there is little material that could properly be called coal and practically no heavy rock, the consequences being that previous attempts to clean the middling have not at all been successful." Referring again to Table I, Plants 72, 73, 74, and 75, these data were obtained by Yancey and Geer and others and presented at the February 1950 Meeting, AIME, in a paper entitled Laboratory Performance Tests of the Humphreys Spiral as a Cleaner of Fine Coal. Results shown for those tests involve all particles from 8 mesh through the colloids, which admittedly is not an ideal situation for spiral feed if much refuse is contained in the —80 mesh or —100 mesh size range. As an illustration of the effect of treating too broad a size range, let us consider Plant No. 75, Kentucky No. 9 seam. Spiral feed was 8 mesh x 0. Now had this been 8x100 mesh the percentage of misplaced material would have been 8.0 pct instead of the reported 15.26 pct. Similar comparisons can be made on the other data presented with respect to the spiral. Other types of equipment show a similar trend in that whenever too fine a size is treated in a given unit process the percentage of misplaced material increases. Since the spiral is working near the finer end of the size range, it will sometimes be advantageous to treat the entire range of —8 mesh material rather than to deslime and make a fair showing on, say, the +80 or + I00 mesh. Desliming is subsequently done in any case in the dewatering or thickening operation. Results obtained by spiraling any given coal depend on factors too numerous and complex to discuss here, but there are strong indications that proper preparation of feed to the spiral can improve results obtained on some of the raw coals tested. This is clearly pointed out at the end of the aforementioned paper by Yancey and Geer. "The spiral is an extremely simple device which involves no moving parts and is constructed almost entirely of unmachined castings. Since it is such an uncomplicated mechanism, operation is simple and virtually foolproof. These characteristics, which go far toward insuring low cost operation, are attractive attributes in any coal cleaning unit." Certain equipment used in conjunction with the spiral has resulted in a decrease in the percentage of misplaced material, notably in actual practice the launder screen which is used to remove objectionable high ash fines from a spiral-washed coal product. Private correspondence with the U. S. Bureau of Mines has intimated that an additional yield of coal is possible by flotation of the spiral middling. Possible future improvements and developments may result from other methods now under consideration. Finally, some compromise must be made between the best metallurgical performance and the best practical or economical results. Mr. Lyons emphasizes in his summary this objective of overall economy in selection of equipment. G. B. Walker (American Cyanamid Co., Stamford, Conn.)—I had the pleasure of reviewing the draft of this paper and my curiosity was aroused by the data given for Tromp plants, in that all of the examples shown appeared to be 2-product separations, whereas all the Tromp plants with which I am familiar have been 3-product units. The data given for plants No. 101 and 102 appear to be taken from Tromp&apos;s brochure on his process and represent the results obtained at the Dominale plant in Holland which was operated for many years by Mr. Tromp. The plant, which was designed to treat 58 tons per hr, was sampled while treating 35 tons per hr of 3Y4x-in. coal. Example 101 appears to conform to what would result if the middling product were calculated into the refuse product, while Example 102 represents the calculation of the middling into the coal product. It is believed that Examples 103 and 104 represent the operation of the Willem-Sophie Mine in Holland recalculated on the same basis. In checking the English examples given by numbers 14, 15, 16, and 17, the same procedure seems to have been followed. These results have, apparently, been taken from an article in Colliery Engineering in August 1941, describing the initial operation of the Williamthorpe Colliery of the Hardwick Colliery Co. Two vessels are employed in this plant, one to treat soft coal and one to treat hard coal. Example No. 14 presents the results that could be obtained from the soft coal bath if the middling were calculated into the refuse, and Example 15 the results when the middling is calculated into the coal. Examples 16 and 17 represent the same expedient in the case of the hard coal bath. Of interest to this discussion is the fact that during the past year the Simon-Carves Engineering Co. in England has installed in the Williamthorpe plant their new "Sim-Car" medium cleaning system which is based on magnetic extraction and control and which is licensed under the Heavy-Media Separation Processes patents by the American Zinc, Lead and Smelting Co. This system has been described in the December 1951, issue of Colliery Engineering. It is reported that since the Williamthorpe Colliery was changed from the Tromp system of medium cleaning to the Magneto-Motive method of medium control the opera-

Jan 1, 1953

By Orville R. Lyons

DISCUSSION Judson S. Hubbard (The Humphreys Investment Co., Denver)—In this very interesting paper several brief references are made to the Humphreys spiral, a device used for cleaning fine coal. In Table I, Plants 76 and 77, data are given on spiral performance treating Raton and Trinidad coal. The fine coal, as fed to the spiral in these instances, is actually a table middling, hence the more easily treated material was previously removed and a large number of particles were present which were difficult to clean. Mr. W. M. Bertholf of the Colorado Fuel and Iron Co. presented a paper in February, 1946, Cleaning Table Middlings from a Coal Washery with the Humphreys Spiral Concentrator, from which I quote: "In considering the results of our tests it should be noted that the feed was our table middling, and that any real separation is a &apos;moral victory&apos; as there is little material that could properly be called coal and practically no heavy rock, the consequences being that previous attempts to clean the middling have not at all been successful." Referring again to Table I, Plants 72, 73, 74, and 75, these data were obtained by Yancey and Geer and others and presented at the February 1950 Meeting, AIME, in a paper entitled Laboratory Performance Tests of the Humphreys Spiral as a Cleaner of Fine Coal. Results shown for those tests involve all particles from 8 mesh through the colloids, which admittedly is not an ideal situation for spiral feed if much refuse is contained in the —80 mesh or —100 mesh size range. As an illustration of the effect of treating too broad a size range, let us consider Plant No. 75, Kentucky No. 9 seam. Spiral feed was 8 mesh x 0. Now had this been 8x100 mesh the percentage of misplaced material would have been 8.0 pct instead of the reported 15.26 pct. Similar comparisons can be made on the other data presented with respect to the spiral. Other types of equipment show a similar trend in that whenever too fine a size is treated in a given unit process the percentage of misplaced material increases. Since the spiral is working near the finer end of the size range, it will sometimes be advantageous to treat the entire range of —8 mesh material rather than to deslime and make a fair showing on, say, the +80 or + I00 mesh. Desliming is subsequently done in any case in the dewatering or thickening operation. Results obtained by spiraling any given coal depend on factors too numerous and complex to discuss here, but there are strong indications that proper preparation of feed to the spiral can improve results obtained on some of the raw coals tested. This is clearly pointed out at the end of the aforementioned paper by Yancey and Geer. "The spiral is an extremely simple device which involves no moving parts and is constructed almost entirely of unmachined castings. Since it is such an uncomplicated mechanism, operation is simple and virtually foolproof. These characteristics, which go far toward insuring low cost operation, are attractive attributes in any coal cleaning unit." Certain equipment used in conjunction with the spiral has resulted in a decrease in the percentage of misplaced material, notably in actual practice the launder screen which is used to remove objectionable high ash fines from a spiral-washed coal product. Private correspondence with the U. S. Bureau of Mines has intimated that an additional yield of coal is possible by flotation of the spiral middling. Possible future improvements and developments may result from other methods now under consideration. Finally, some compromise must be made between the best metallurgical performance and the best practical or economical results. Mr. Lyons emphasizes in his summary this objective of overall economy in selection of equipment. G. B. Walker (American Cyanamid Co., Stamford, Conn.)—I had the pleasure of reviewing the draft of this paper and my curiosity was aroused by the data given for Tromp plants, in that all of the examples shown appeared to be 2-product separations, whereas all the Tromp plants with which I am familiar have been 3-product units. The data given for plants No. 101 and 102 appear to be taken from Tromp&apos;s brochure on his process and represent the results obtained at the Dominale plant in Holland which was operated for many years by Mr. Tromp. The plant, which was designed to treat 58 tons per hr, was sampled while treating 35 tons per hr of 3Y4x-in. coal. Example 101 appears to conform to what would result if the middling product were calculated into the refuse product, while Example 102 represents the calculation of the middling into the coal product. It is believed that Examples 103 and 104 represent the operation of the Willem-Sophie Mine in Holland recalculated on the same basis. In checking the English examples given by numbers 14, 15, 16, and 17, the same procedure seems to have been followed. These results have, apparently, been taken from an article in Colliery Engineering in August 1941, describing the initial operation of the Williamthorpe Colliery of the Hardwick Colliery Co. Two vessels are employed in this plant, one to treat soft coal and one to treat hard coal. Example No. 14 presents the results that could be obtained from the soft coal bath if the middling were calculated into the refuse, and Example 15 the results when the middling is calculated into the coal. Examples 16 and 17 represent the same expedient in the case of the hard coal bath. Of interest to this discussion is the fact that during the past year the Simon-Carves Engineering Co. in England has installed in the Williamthorpe plant their new "Sim-Car" medium cleaning system which is based on magnetic extraction and control and which is licensed under the Heavy-Media Separation Processes patents by the American Zinc, Lead and Smelting Co. This system has been described in the December 1951, issue of Colliery Engineering. It is reported that since the Williamthorpe Colliery was changed from the Tromp system of medium cleaning to the Magneto-Motive method of medium control the opera-

Jan 1, 1953

By S. M. Javad Koleini

Processing of phosphate ore sample was investigated for increasing phosphate content and reduction of its gangue minerals such as carbonates and silicates. Experimental designs were per-formed using Taguchi method. Different flotation methods were investigated for Moundon phosphate ore sample and three methods were used as follow: direct cellophane flotation, reverse calcite flotation and reverse quartz flotation. Because of the sensitivity of cellophane flotation to slime content, prior desliming was carried out before the experiments. Direct flotation experiments carried out and a concentrate contained 23.80% P2O5, 6.88% SiO2 and 72.33% recovery were produced. The achieved P2O5 content in the produced concentrate from reverse flotation of calcite was 26.44% and the recovery was 82.56%. The concentrate products obtained at optimum conditions by direct flotation of phosphate and indirect flotation of calcite were subjected to indirect flotation of calcite and direct flotation of appetite and final concentrates with P2O5 grade of the order of 24.19% and 26.96% were obtained, respectively. It was concluded that one stage direct flotation is unavoidable to achieve a product with low SiO2 content.

Jan 1, 2014

By R. S. Sherwin

The extractive metallurgy of primary aluminum from its ores is discussed with special attention to the production of alumina from high grade ores by the Bayer process, including differences between American and European practice and a brief description of some processes for lower grade ores and the electrolytic reduction of the oxide to aluminum. METALLIC aluminum is not found in nature, but the oxides, hydroxides, and especially the silicates are plentiful. The estimated percentage of in in the crust of the earth is about 8 pct while that of iron is about 5 pct. By far the larger portion of this is combined with silica in the form of various clay minerals and igneous silicate rocks. From the point of view of extractive metallurgy of aluminum, these are low grade ores while the better qualities of bauxite are the high grade ores. There have been various definitions of bauxite but perhaps the best definition at the present time is that bauxite is a rock or earth commonly used as an ore of aluminum or its salts in which the aluminum is present predominantly as a hydrate or a mixture of hydrates and hydrous oxides. It contains varying amounts of oxides of silicon, iron, and titanium and traces of compounds of some of the less common elements. The silica is mainly combined with alumina as clay or clay minerals which are hydrous aluminum silicates, although a part of it may be present as quartz sand. On the American continents, the alumina is mainly present as gibbsite, Al2O3 . 3H2O, and the same may be said of the best known deposits of the Dutch East Indies and some of the deposits in India. In France and other countries in Europe as well as in Africa, the alumina is present mainly as boeh-mite, A12O3 . H2O, but in some of these deposits it is mixed with minor amounts of gibbsite. Some other deposits, such as those in the islands of Haiti and Jamaica, evidently contain two or more hydrates or hydrous oxides of alumina in varying proportions. Perhaps the main portion of the alumina may be present as gibbsite and boehmite with the proportion between the two varying rather widely. In the silicate minerals, including clay, the alumina is chemically combined with silica and has not been separated satisfactorily by mechanical or physical ore-dressing methods. Low grade bauxites are mixtures of hydrates, usually gibbsite or boehmite, with clay, iron oxides, etc. In some low grade bauxites, it is possible to separate a portion of the gibbsite or boehmite, which may be present as relatively hard and coarse particles, from soft or finely divided clay minerals by log washing or similar methods. This has been applied to the product of some mines or parts of them, but on other ores it is not applicable. In some cases the gibbsite or boehmite is almost as fine and soft and of nearly the same specific gravity as the clay minerals so that washing and gravity separations are not successful. The iron oxide, the clay minerals, and a part of the titanium minerals are often so finely dispersed in the ore that any of the physical mineral separation methods, including separations by gravity, particle size, flotation, and electrostatic or magnetic separation, have not been commercially SUCCESSFUL except on relatively small lots of ore. For these reasons, the only available methods of separation on the general run of ores have been methods which would be classed as chemical rather than physical or mechanical separations. Aluminum oxide can be reduced by carbon at temperatures above 1800°C to form metallic aluminum and aluminum carbide or nitride. The temperature for rapid reduction of aluminum oxide to metallic aluminum is about the boiling point of aluminum and above the temperatures necessary to reduce iron, silicon, and titanium so that the direct reduction of an aluminum ore with carbon will produce an alloy of aluminum, iron, titanium, silicon, etc., which may be mixed with carbides and nitrides. Also a large amount of the reduced aluminum may be lost as a vapor except in the presence of some alloying agent such as copper or other metals. While it is possible to refine such alloys or mixtures so as to produce commercially pure aluminum, the methods which have been found are too expensive for the present market. One direct reduction method which found limited commercial use in Germany during World War II was the direct reduction of a mixture of clay containing very little

Jan 1, 1951

By Ivor Jenkins

PROCESSES employing controlled atmospheres in the heat-treatment of metals and alloys are now well established on an industrial scale, and the general principles involved and the advantages to be gained in the use of the technique have received considerable publicity over a period of years. During this time the greatest emphasis has been on the establishment of the process in industry, the main effort being devoted to development work on various types of atmospheres, their useful fields of application and the design of suitable heat-treatment furnaces in which they are to be used. The various types of controlled atmospheres are fairly well standardized, but there still remains a considerable field for the investigation of related problems, in connection with both the gas-generating equipment and the surface chemistry of the heat-treatment of metals, especially in complex controlled atmospheres. This paper describes certain investigations of this nature that were carried out in England up to 1939. The development of processes employing controlled atmospheres has been associated to a very large extent with the heat-treatment of steels, both because of the major importance of steel in industry and the many diverse and difficult problems of surface chemistry involved when steels of such a wide range of composition are heated in the presence of active gases Controlled atmospheres such as dissociated and burnt ammonia or charcoal gas are used in the heat-treatment of low-carbon and high-carbon steels, but from certain points of view a suitable atmosphere generated by the partial combustion of city gas is to be preferred. Both the ammonia derivatives are comparatively expensive, and even if the burnt ammonia is regenerated, thereby reducing operating costs, the capital expenditure of the gas plant may make it prohibitive for the smaller type of furnace installation. Charcoal-gas generators are not without their operating difficulties: especially those associated with the quality of the charcoal, clinker formation and the removal of carbon dust from the generated gas. City gas has the advantage of being comparatively cheap and "on tap" in most if not all industrial centers. The use of raw city gas as a protective atmosphere in the heat-treatment of steels has been described,1 but it presents certain difficulties because of its complex composition and its explosive nature when mixed with air. These factors, together with the tendency toward sooting at elevated temperatures, arising mainly from the hydrocarbon gases present, all militate against its use as a protective atmosphere in industrial furnaces. The following analysis, based on the analyses of city gas in different localities in Great Britain, indicates the wide limits existing

Jan 1, 1947

By Henry T. Miller

INTRODUCTION The basic purpose of this paper is to suggest procedures and practices to insure that no employee working in a uranium milling operation receives exposure to radioactive, toxic, or other materials or agents that might produce a permanent, deleterious effect upon his physical health and well-being. This program is also designed to insure that each employee can carry out his assigned duties without risk to his health or to that of his fellow workers. The assumptions underlying this program are by necessity, conservative, with emphasis being placed upon prevention rather than correction or reaction to an undesirable situation. The limits specified-either as Threshold Limit Values (TLVs), Maximum Permissible Concentrations (MPCs), or Radiation Protection Guides (RPGs)--represent both standards of best practice and those of law and regulation. The total program is envisioned as a balanced combination of occupational hygiene and radiation monitoring. This includes surveying, air sampling, personnel dosimetry, bioassay, medical surveillance, epidemiology, and training--all backed by a thoroughly tested and evaluated set of emergency procedures. The program, as presented, is keyed to the results of monitoring, surveying, air sampling, medical surveillance, and epidemiology--it being obvious that no problem can result when no hazard can be identified. The lack of an identified hazard, however, does not negate the need for such a program and prudence suggests the need for routine surveillance programs in the best-designed facility. [The Model Mill] The model mill used in this paper is that presented by the U. S. Environmental Protection Agency in its study of the uranium fuel cycle1. The model mill processes 600,000 metric tons of ore per year into 1,140 metric tons of U308 as yellowcake per year using an acid leach process. Other characteristics of the mill such as location, tailings dam design, number of light water reactors supported, and geographical location are not pertinent to this paper. The usual remote western location of these facilities produce some difficulties in delivering the required occupational medical services and is not within the scope of this paper. The operating lifetime of most mills is commonly estimated as from 12 to 15 years in actual practice; however, the U. S. Environmental Protection Agency estimated a 20-year lifetime for its model. The principal steps in the acid leach process include :2 (a) Ore is blended and crushed to pass through a 2.5 cm (one inch) screen. The crushed ore is then wet ground in a rod or ball mill and is transferred as a slurry to leaching tanks. (b) The ore is contacted with sulfuric acid solution and an oxidizing reagent to leach uranium from the ore. The product liquor is pumped to the solvent-extraction circuit while the washed residues (tailings) are sent to the tailings pond or pile. (c) Solvent extraction is used to purify and concentrate the uranium. (d) The uranium is precipitated with ammonia and transferred as a slurry. (e) Thickening and centrifuging are used to separate the uranium concentrate from residual liquids. (f) The concentrate is calcined and pulverized. (g) The concentrate or yellowcake is packaged in 208 liter drums for shipment. This process is illustrated in Figure 1.3

Jan 1, 1981

By C. L. Dake

AN insistent and steadily growing demand is evident for the broadening of undergraduate curricula in engineering. Among suggested additions are training in public speaking, report writing, business law and contracts, accounting, finance, money and banking, scientific management, economics, social and labor problems, psychology, and ethics. The demand for these subjects comes from the older engineers and from the industries that employ our graduates. Of course, many of the men who suggest such changes are influenced largely by their personal needs and contacts, but the demand is sufficiently widespread to merit serious study by educators. In fact, many educators themselves concur in the proposals for broader training. These matters were examined intensively a few years ago in an exhaustive study of engineering education. Opinion at that time was rather strongly against any profound change. Since then, however, the whole economic system has undergone violent disruption. Even if recommendations made a few years ago were appropriate to the conditions then- obtaining, the enormously changed conditions of the present make advisable a further careful study of this problem. In recent years, mergers, to reduce costs, have sharply limited the number of engineers employed and have thereby intensified the struggle to secure positions. In this stiffer competition, engineers with the most adequate training undoubtedly fare the best.

Jan 1, 1934

By Frederick C. Trutt

Research performed on arcing faults and their interactions with electrical insulating materials within coal-mine power system enclosures are described. This work was performed at The Pennsylvania State University from December 1, 1976 through August 31, 1978. The thrust of this research has been divided into several primary areas. These include the development of a theoretical procedure for computation of three-phase and line-to-line arcing fault currents as a function of time, formulation of arc and tracking models so that arc voltages and energies may be computed, development of a simulation for pressure and temperature rises when a fault interacts with an electrical insulating material within an enclosure, experimental evaluation of electrical insulating materials in terms of their tendency to support and maintain an arc as well as their potential contribution to enclosure pressure rises, and integration of the above research areas to form overall procedures for theoretical and experimental evaluation. During the course of this research, consideration has been given to the possible effects? of an internal methane-air explosion within an enclosure as well as the potential influence of utilization-voltage-level, conductor spacings, and enclosure venting. Conclusions and recommendations address potential means for protection against enclosure failure, internal design considerations, and low-power testing procedures which might be used to evaluate insulating materials to be utilized within an enclosure.

Jan 1, 1978

By Meghan M. Murphy, Derek Penrice

The Downtown Rail Extension (DTX) is principal element of the planned second phase of the Transbay Joint Powers Authority’s (TJPA) Transbay Program. When complete, this regionally critical project will connect Caltrain’s regional rail system and the California High-Speed Rail Authority’s statewide system to the Salesforce Transit Center in downtown San Francisco via a 1.3-mile tunnel. Studied since the 1970s, the DTX has now received environmental clearance, and key stakeholders are aligned on the project’s configuration and its expedited delivery. This paper describes efforts to optimize the DTX configuration and develop an informed project delivery strategy that incorporates construction industry feedback on project phasing, contract packaging, and procurement methods that accomplish the project’s budget goals.

Jun 13, 2021