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Discussion of the paper of Frank H. Kneeland, Safeguarding the Use of Mining Machinery. By B. F. Tillson and Arthur Williams. 1073 Discussion of the paper of Howard N. Eavenson, Safety Methods and Organization of United States Coal & Coke Co. By William H. Grady, Thomas W. Dawson, C. W. Goodale, J. Parke Charming, B. F. Tillson, George S. Rice, Arthur Hovey Storrs, Harry H. Stoek, Lawson Blenkinsopp, Arthur Williams, and Howard N. Eavenson. 1075 Discussion of the paper of J. Parke Channing, Enlarging the Worth of the Worker and the Perspective of the Employer. By Fred H. Rindge, Jr., S. A. Taylor, Elmer L. Corthell, and Philip W. Henry. . 1095 Discussion of the paper of Walther Mathesius, High Blast Heats in Mesaba Practice. By Joseph W. Richards, W. A. Forbes, J. E. Johnson, Jr., H. P. Howland, and Walther Mathesius. . 1100 Discussion of the paper of W. R. Shinier, Effect of Finishing Temperatures of Rails on Their Physical Properties and Microstructure. By William R. Webster, Henry M. Howe, A. A. Stevenson, G. K. Burgess, and Samuel L. Hoyt . 1107 Discussion of the paper of George K. Burgess and Sir Robert A. Hadfield, Sound Steel Ingots and Rails. By Albert Sauveur, Henry M. Howe, Henry D. Hibbard, Joseph W. Richards, F. W. Wood, and George K. Burgess. . 1112 Discussion of the paper of Henry M. Howe and Arthur G. Levy, Are the Deformation Lines in Manganese Steel Twins or Slip Bands? By J. E. Stead 1116 Discussion of the paper of F. C. Langenberg and R. G. Webber, Structures and Hysteresis Loss in Medium-Carbon Steel. By Henry M. Howe. . . 1117 Discussion of the paper of Charles H. MacDowell, German and Other Sources of Potash Supply. By George S. Rice, Charles H. MacDowell, and V.C. Grubnau . . 1118 Discussion of the paper of Joel H. Watkins, White-Burning Clays of the Southern Appalachian States. By H. Ries 1120

Jan 5, 1915

By U. N. Bhrany, N. Arbiter

F. C. Bond (Consulting Engineer, Processing Machin-Dept., AllisChalmers Manufacturing Co., Milwaukee) This is a very comprehensive paper. It deals with 1) size distribution functions, 2) energy-particle size relationships, and 3) criteria for grinding efficiency. The method of analysis used consists of tracing the increase and decrease of successive Tyler scale screen fractions as grinding progresses. The treatment is based upon the assumption that the work done in grinding is concentrated entirely upon the coarse size fractions in the mill. It is postulated that when a feed consisting entirely of particles of one size fraction is batch ground that finer fractions increase at a constant rate as grinding progresses, until finally each in turn begins to be ground and to decrease in amount. It is shown that under these conditions the empirical Gates-Gaudin-Schuhmann log-log size distribution plot can be logically explained. It is difficult to imagine any condition of ball mill operation, or any probability analysis, which would result in no grinding contacts on the finer particles present. However, the authors' postulate must be accepted if their conclusions are to be sustained. They conclude that the equation expressing the for mation rate of each size fraction can be resolved into the G.-G.-S. size distribution equation, and that the curvilinear portion of the log-logsize distribution line represents the only particle sizes which are being ground as well as produced. The author's statement that the G.-G.-S. Eq. 4 applies only to the finer portion of the size distribution while the Rosin-Rammler Eq. 6 applies over the whole size distribution range certainly supports the view that an exponential equation of the Rosin-Rammler type corresponds more closely to actual size distribution than does the power law G.-G.-S. equation. However, the G.-G.-S. equation is used to derive Eq. 8 and Eq. 9, which state that the energy required to grind a ton through a particular mesh varies inversely as the particle size to the exponent -a , whic h is the slope of the extended G.-G.-S straight line. This is stated to apply to both batch and continuous grinding. The G.-G.-S. slope exponent thus becomes very important in all crushing and grinding analyses by this method. The authors include some data which can be used to test their conclusions, as described below: We will consider grinding a feed composed of 14 mesh (1190u) to -200 mesh (74u). The average of 17 Allis-Chalmers grindability tests on quartz gives a work index (Wi) of 12.77. The -200 mesh product should have 80 pct passing 41.6u or P = 41.6. The equivalent 80 pct passing size of the 14 mesh feed particles is considered to be the particle diameter, as is true for spherical particles. Thus F is 1190u. In the work index equation 10 wi 10 wi 127.7 127.7 P f 41.6 1190 = 16.1 kw-hrperton. In Table I the ball mill exponent for quartz is 1.0 and from Eqs. 9 and 10: w=kw-hr/ton =Constan/(do/di) = 16.1 ton so that (1190) Constant/ Constant and the Constant equals 259. The tons per horsepower-hour value is 0.0464, in agreement with Table IV. The exponents for other minerals are listed in Tabl e 11. In Table IIA, the kilowatt hours per ton required to grind these minerals between the same sizes as the quartz, both by the authors' method and the work index equation, are listed. According to the authors kw-hrperton = 259/(16.1)".

Jan 1, 1961

By S. R. B. Cooke, Iwao Iwasaki, Hyung Sup Choi

John Dasher (Materials and Processes Engineer, Crucible Steel Co. of America, Pittsburgh, Pa.) The authors are to be congratulated for their excellent work, dramatic results, and persuasive explanation on the neglected subject of the temperature variable in soap flotation. The explanation includes why heat is helpful in conditioning and why heat is helpful in flotation. This is a fine start, but I hope the work will be continued to show which is the major effect. In the work of Noman and Dash er(1) it was found that substantially all the benefit of heat came from the conditioning step, and that results were much better if a thick pulp was conditioned hot and then diluted with unheated (or less heated) water for flotation (or agglomerate tabling) than if the same quantity of heat was used to condition and float at some intermediate temperature. In other words, heat for conditioning gives a Rood return, but additional heat for dilution water is near or beyond the point of diminishing return. Strathmore R. B. Cooke, lwao lawsaki, and Hyung Su0p Choi (Authors' Reply) Mr. Dasher'skind comments on our paper are appreciated. In view of the relative simplicity of empirical investigationsof the temperature variable upon conditioning and flotation, it ispuzzling why the subject is so neglected. Our work, being sub sidiary to the main investigation of the effect of structure of fatty acid collectorsupon iron ore flotation, was quite restricted, but we intend to return to the subject as soon as possible. Mr. Dasher's point, that given a certain number of calories it is better to invest them for conditioning in a thick pulp rather than to spread them out for both conditioning and flotation operations, is well taken and cannot be argued. On the other hand there is no question that high temperature during the flotation of iron ores induces superior frothing characteristics, and, in particular, greatly augments the rate of flotation and markedly decreases the time required for "difficult" particles to float.

Jan 1, 1961

By R. Schuhmann

D. W. Fuerstenau (Associate Professor of Metallurgy, Dept. of Mineral Technology, University of California, Berkeley, Calif.) In his excellent paper, Dr. Schuhmann has proposed a mechanism for conlrninution which can be utilized to explain a number of phenomena which have been observed in the size reduction of solids. The difference between the product obtained from a ball mill and from a rod mill can readily be explained by Dr. Schuhmann's hypothesis. Consider that the size distribution of the product from each comminution event is given by y; - 100 (x/ki) (1) where Yi is the cumulative percent finer than size x, ki is the size modulus, and is the distribution modulus for the single comminution event. The total product from a cornminution operation will be the summation of the products from all of the individual cornrninution events. If the size moduli from the various comminution events range from kl to kn , the size distribution of the total product will be a straight line on a log-log plot only to a particle size equal to k ,. Considering 100 volurnes of feed, the value of y for x equal to kl is Riven by where y is the cumulative percent finer than size kj, z is the number of comminution events of kind i, and vi is the volume of each particle crushed. y=z1 qv1+z2 qv2 (k1/k2) +.............zn qvn (k1/kn) (3) F or particle sizes greater than k1 7 e. 6. x = k3, the value of y will be given by k3 A y = z1 qv1 + z2bv2 + Z3uV3 + zn zn qvn (k3/kn)(4) Only when x is equal to k, will y equal 100. 'Thus, the total product from a comminution operation will be a straight line on a log-log plot strictly up to a particle size equal to kl and above that size the line will curve so that y will equal 100 for the coarsest particle or k, (see Fig. 1). The actual shape of the size of distribution curve will depend upon the relative number of conmninution events of different kinds. In the case of ball milling, the ratio kn/kl may be pite large and there will be considerable deviation at the coarse sizes in a Schuhmann plot. kIowever, the grinding action in a rod mill is such that fine particles are protected from the rods by the coarse particles. Consequently, kn/kl will be quite small and the product from a rod mill will give a straight line Schuhmann plot almost to y equal 100%. In his paper, Dr. Schuhmann also has presented a very straightforward definition of grindability derived from the basic energy equation E = Ak-a (5) where E is the energy required to comminute a material to a size modulus of k and A is a constant. According to Dr. Schuhmann, the grindability of a material is the volume of material comminuted finer than unit size per unit of energy expended, this being 1/A. For comparison of grindabilities at any size two numbers are required, both = and 1/A, and anyone tabulating grindabilities should give both of these numbers. Concerning the experimental determination of grindabilities, if the material being comminuted possesses distinct cleavage, the energy re quired for comminution is given by E = Ak-(a-y) (6) where y is small number less than = . Because of cleavage, less energy will be required for comminution. To be certain that cleavage does not affect the gindability of the material being tested, it is suggested that the method used by Dr. Charles for determining grindability be used to ascertain if the deviation exponent y is zero or finite. R. Schuhmann, Jr. (Author's Reply) Dr. Fuerstenau's interesting comparison of rod milling and ball milling

Jan 1, 1961

By J. L. Gillson

Joseph H. Birman (Chairman, Dept. of Geolcgy, Occidental College, Los Angeles, Calif.) Many thanks to J. L. Gillson for so comprehensive a survey of the titanium sand deposits of the world. Over the past two years, the writer studied rather carefully some of the Cretaceous ilmenite deposits in New Mexico and Colorado, including the Sanostee deposit shown in Gillson' s map, Fig. 6. The deposits were geologically mapped and sampled, surveyed by magnetometer, and the Sanostee deposit was drilled. Moat of the genetic interpretations made by Gillson concerning the recent sedimentary ilmenites can be clearly read in the Cre-taceous deposits in New Mexico. The deposits are in marine regressive members of the Mesaverde group. The deposits trend within a few degrees of N 55° W and persistence of this trend through both distance and time (the deposits are in different members of the Mesaverde group) implies that the deposits are offshore bar accumulations. Heavy mineral content ranges from about 10 pct to more than 50 pet. Thicknesses are from 4 ft to as much as 20 ft, widths range from 150 to 600 ft, and lengths range from a few thousand feet to a few miles. All are locally or completely overlain by carbonaceous shale and thin coal. The deposits are strongly altered, and from X-ray diffraction the principal titanium alteration product appears to be anatase. Drilling of the Sanostee deposit shows that the alteration is not related to the present erosion surface but took place at the site of deposition and before burial by the overlying sediments. Iron oxide derived from the alteration cements the grains and is the chief coloring agent in the immediately overlying shales. In the Sanostee deposit, there is virtually no feldspar, the gangue mineral being almost exclusively quartz. Sorting is excellent, but the quartz grains are subangular to angular. Petrographic examination reveals that the angularity is not due to overgrowths. Therefore, short distance transport is implied. The mineralogical composition suggests a low relief granitic terrane as the source area. Feldspar may have been decomposed by weathering at the source, but an alternative preferred by the writer is that each of the deposits represents a long-term concentration site (for example a large sand bar) which was continually being deprived of light minerals by wave action and selective transport. Feldspar may have been mechanically reduced in size more quickly than the quartz, and thus its rate of removal may have exceeded that of quartz. Concentration of heavy minerals would thus occur until the marine regression removed the zone of maximum wave energy from the site. Chemical alteration would then become the most important process until burial by encroaching swamp deposits.

Jan 1, 1961

By R. F. Shurtz

R. Duval (Mining Engineer, Ancien eleve de PEcole Polytechnique, Paris, France) I do not agree with the Eq. 3, reading: m =1/100- [(0.214x30.4) + (0.7B6 x0.00)] =6.5pct CaO If 0.214 and0.786 were proportions by weight, the equation would represent the well known mixtures law of the conventional arithmetics and 6. 5 pct CaO would be the correct average content. But it is not the case as the author states: "In samples consisting of single grains of mineral, those grains must, as already mentioned, be either of dolomite or magnesite. Since 78.6 pct of the deposit consists of magnesite and 21.4 pct of dolomite (excluding for present purpose the presence of other minerals), for any single grains picked at random the probability will be 0.214 that is it dolomite and0.786 that is it magnesite. In 1000 such samples the expected numbers of dolomite and mapesite grains will be 214 and 786 respectively." 0.214 and 0.786 would be proportions by weigbt under the necessary condition that all grains of dole mite and magnesite should have an identical weight. Obviously it is not the case, as the specific gravities are not the same for mapesite and dolomite and the volumes of the grains are different. Furthermore, because of these differences the conditions for a random sampling are not fulfilled and we are not authorized to state that the probabilities are, respectively, 0.214 and 0.786. The author however makes a simple application of Eq. 1: M = 1/n— ? fi x i . n Should we deduce that this relation is wrohg? Not at all, but when applying Eq. 1 you must not overlook what it actually. means. Eq. 1 gives a definition of the arithmetic mean of a total of n observed values Xi and nothing else. But the average conteht of a deposit has not the same significance. It is the ratio between the weight of concerned mineral in the deposit and the total weight of the deposit. As from 1000 particles the 214 of dolomite and the 786 of magnesite have not the same weight, the two definitions do not concur, and when applying Eq. 1 the result is an arithmetic mean of figures which has no connection with what is named average contentof a deposit. The situation is similar to the calculation of an average velocity. If a car travels a first mile over at 30 miles per hr and a second mile over at 60 miles per hr, when applying formula 1 you find as average velocity for the 2 miles: 30+60 ------- - 45 miles per hour. Many people calculate in this way and they do not realize that a mistake is involved. In fact the definition of he average velocity for the 2 miles is the quotient of the distance of 2 miles by the time (in hours) necessary for 2 miles travel, i.e.: 2 ---------- = 40 miles per hr. 1 + 1 30 60 In other words, the average volocity wanted is not the arithmetic but the harmonic average of the two velocities. The above mentioned bias in the calculation of the average contents of deposits is frequent, even in the assessments made by experienced engineers and is independant of what is named the sampling error. In order to supress the bias and to be able to use Eq. 1, you must apply a correction. An example on the subject can be found in an article by Duval et al. in the January 1955 issue of the ''Annales des Mines" (French), page 19. R. F. Schurtz (Author's Reply) Mr. Duval's position is quite correct. The proportions shown for dolomite and magnesite., respectively, of 0.214 and 0.786 are, in fact, proportions by weight uncorrected for specific gravity. In our day to day operation of producing magnesite from these mines at a very substantial rate, we do not normally make corrections for the difference between the specific gravity of dolomite and that of magnesite. If these corrections are made in Eq. 3 as shown in my article, then the numbers of grains turn out to be in proportions of 0.226 dolomite and 0.774 mapesite instead of the values actually shown in the equation. For the purposes of our work, and in view of the inherently low accuracy of the data, this correction was not deemed worthwhile making.

Jan 1, 1961

By Garth M. Crosby

Hugh E. McKinstry (Professor of Geology, Laboratory of Mining Geology, Harvard University, Cambridge, Mass.) Pre-Cambrian age for the Coeur d'Alene mineralization, as indicated by isotope studies and recently discussed by Silverman, Long, and Kulp' would explain some features of the mineralogy of the district that have long been a puzzle. In the Silver Belt, especially in the Sunshine mine, lead-zinc min erali zation passes downward into tetrahedrite ore. Although commercially speaking this is a rich silver zone, the silver is, quantitatively, a mere impurity amounting to a small fraction of 1 pct. In effect the zonal change is from lead-zinc down into copper mineralization—-the usual zonal change that occurs in many districts. In the Burke-Mullan portion of the district, on the other hand, lead-zinc ore does not pass downward into copper mineralization of the Sunshine type. Instead, pyrrhotite appears in depth while the tenor in lead and zinc declines. The change occurs at increasingly greater depth away from the Gem stocks, deepening irregularly by some 3000 ft within two to four miles. As noted by Crosby, biotite, grunerite, and garnet appear in the deep zone. The marked difference between this apparent zonal change and the more nomal change to a copper zone is quite explainable if already existing pre-Cambrian veins have been locally metamorphosed by a Tertiary intrusive. The chief obstacle to this explanation is the evidence that mineralization cuts Laramide intrusives, but evidence on this point is conflicting as indicated by Crosby's observations and summary. This is not the first instance of a problem in determining the time-relationship of mineralization to a spatially associated intrusive. At the Noranda mine veins of chalcopyrite extend into a diabase dike which apparently cuts the ore, although Price2 concludes that the ore is later. However, at the Normetal mine diabase of a similar dike is thought3 to intrude the orebody but is cut by stringers of chalcopyrite. In the Sullivan Mine, B. C.,4 galena follows a fracture in a larnprophyre dike which, itself, cuts the main orebody. In the Sudbury district dikes of olivine diabase are clearly younger than the nickel ore, yet where they cut the Creighton orebody they contain stringers of chalcopyrite. In the laboratory, Hawley5 has developed veinlets of sulfide (chalcopyrite) it] silicate gangue by heating specimens of ore in a stream of H2S at 410 °C. From this and other evidernce. it seems clear that sulfides can be transferred from one site to another — at least through short distances — at temperatures well below their melting points. It would seem then that even if sulfide Mineralization in the Gem stocks can be shown unequ ivocably to be later than the intrusive the possibility of "remobiliza tion" need not be more than local. Thus the zonal relationships are quite in accord with a pre-Cambrian age for original deposition of the lead, zinc, copper, silver, and uranium in the Coeur d'Alene district. REFERENCES 1. A. Silverman, A. Long, and J. L. Kulp: Discussion of: The Gem Stocks and Adjacent Orebodies, Coeur d'Alene District, Idaho. Mining Engineering, May 1960, pp. 470-471; AME Trans., 1960, vol. 217, p. 117. 2. Peter Price: Geol. & O. D. at the Home Mine, Noranda, Quebec. Can. Inst. M & M Bull 263, 1934, p. 108; also,Trans. CIMM 1939, pp. 108-140. 3. w. L. Brown Nometal Mine in structural Geology of Canadian Ore Deposits, Montreal, 1948, p. 690. 4. C. O. Swanson and H. C. Gunning: Geology of the Sullivan Mine. Can. Inst. Mining & Met. Trans. vol. 48, 1945, p. 653. 5. J. E. Hawley: Heat Effects on Sulphides and Possible Applications. Univ. Toronto Studies, Geol. Ser. No. 46, 1941, p. 34.

Jan 1, 1961

By Rudolph G. Wuerker

Charles T. Holland (Head, Dept. of Mining Engineeri*, Virginia Polytechnical Inst., Blacksburg, Va.) Mr. Wuerker has presented a very interesting discussion of the use of triaxial test methods for investigating the strength properties of rocks. Such methods, no doubt, eventually will develop considerable information of interest to those concerned with the design of mine layouts, particularly in the field of pillar design. From his discussion of my recent article, "Cause and Occurrence of Coal Mine Bumps" (Holland Mining Engineering 1958, p. 933-1002), it is evident that in one place at least I did not make my meaning clear to him and perhaps others. To clear the matter up I think it best to quote from the article, somewhat more fully than did Mr. Wuerker, as follows: "4) In actual operations — because rocksare not perfectly elastic, homogeneous, nor isotropic and because local yield does occur — the maximum stress as demonstrated by Phillips (Ref. 22, pp. 64, 65) and indicated by much experience in mining, does not occur at the walls of the opening but at a short distance inside the pillar. Furthermore, the maximum stress does not reach as great a value as theoretical considerations and laboratory experimental methods indicate.* Actual distance inside the pillar, measured from the wall, at which the maximum stress exists, has not been determined. Observations in many mines, however, indicate that this distance could have a mini-value of one to six or eight times the bed thickness and that it is probably affected by width and height of the opening, depth of cover, and relative values of the elasticity and plasticity of materials comprising the roof, floor, and coal seam. The actual value of the stress produced probably lies between the theoretical maximum and the average stress concentration that would be produced if the weight of the strata above the unsupported opening were evenly distributed over the pillars for a distance equal to the opening width." The footnote reference in the above quotation referred to the following: "*For example, the Pocahontas No. 4 coal bed in southern West Virginia is mined under cover up to 1800 ft thick. Development openings are driven 18 to 20 ft wide, and the bed is about 6 ft thick. According to the work of Panek, the tangential wall stress at mid-bed height under these conditions would reach values between 4000 and 5000 psi. Actual tests of 3-in. cubes of this coal show its compressive strength would be much less than this, perhaps as low as 400 psi. Yet the pillars usually show no evidence of failure in these headings. In this same bed at a depth of 800 ft, the author has seen an opening 225 ft between supports lying between two old groves approximately 1100 ft apart. According to the theoretical considerations, the stress in the pillar walls would have been about 18,000 psi, yet the pillar showed little or no evidence of weight. In view of these observations, it is clear that the wall stress does not attain the maximum values indicated by theory." (Underlining added to original wording.) By referring to Fig. 2A of my paper it will be noted that theoretically the maximum pillar stress would occur at the pillar wall, i.e., at the passageway surface of the pillar. Obviously this cannot be correct in the cases of stress ranging from 4000 to 18000 psi since the coal at the surface of the pillar is under no constraint and cannot have a strength much greater than 400 or 500 psi. Hence, my conclusion that the maximum stress does not occur at the wall but back in the pillar some distance from the wall. Since these stresses are pushed back in the pillar from the wall, it is also obvious that the loads transferred to the pillar from the opening will be spread over a greater area and hence Pillar stresses will not rise to the values postulated by theory and photoelastic experiment. Further since to visual inspection the coal along the pillar wall did not appear to be failed the conclusion was reached that the stress shift was caused by local elastic or plastic yield and by difference in the elastic modulus of the rocks composing the mine floor, mine roof, and coal bed. Later on under the heading "Strength of Mine Pillars" (pages 1000-1002) the effects of constraint is briefly described. Also a formula taking into account constraint is developed relating pillar strength to the uniaxial strength of coal and the L/T ratio of the pillar. Since my paper was written, reports of experiments conducted in South Africa (Denkhaus, et. al., 1959), in Sweden (Hast 19581, and in Canada (McInnes, et.al., 1959) reveal that the conclusion expressed relative to the existence of a low stress area existing around the edges of pillars and solid faces as described above is generally correct. But it seems possible that where the wall stress developed is less than the unconfined strength of the rock composing the pillar and where the roof, floor, and pillar

Jan 1, 1961

By AIME

Burton J. Westman-Besides decreasing the diamond size, there appear to be two other approaches open to overcome excessive diamond loss and, more particularly, the rapid diamond polish that took place in this quartz-feldspar rock. First, however, a short discussion of the diamond characteristics should be made in order to discuss the situation more clearly. The West African diamonds which we can assume were used. for this drilling have two disadvantages when drilling hard, exceptionally fine-grained rocks. These are rapid polishing and, if .the bit isn't "drilled in" carefully, a tendency to damage, resulting in excessive diamond loss. The normal wear resistance of any diamond is inversely proportional to the compressive strength of the diamond, therefore only the highest grade, of diamonds free from insipient fractures should be employed. Using upwards of 90 diamonds per carat instead of the20 to 30 per carat not only decreases the total projected area of the cutting points but increases the stress per point which, in turn, causes greater depth of penetration into the rock. Considering this depth of penetration, therefore, suggests an approach other than increasing the, number of diamonds per carat and that is to decrease the actual number of diamond points (also the carat weight per bit) so that the penetrating force on an individual diamond is increased. This penetration increases the chipping or rock shearing action.

Jan 1, 1950

By D. W. Fuerstenau, J. H. Brown

The paper by Professors Sun and Mitchell illustrates the value of ultrasonics as a tool for the mineral engineer. In particular, it opens up a number of possibilities to those interested in the field of comminution. The provision of desliming technique, which is thorough as well as rapid, is a distinct contribution, but the data of the paper suggest applications beyond desliming, in that the same treatments may be useful for size determination purposes. Fig. 1 shows that a marked dependence exists between particle size and the height to which it can be elevated by ultrasonic vibrations. While the curves are only extended down to a size of about 100 u, the fact that they appear to be increasing in slope in this range suggests that the separation may be even more marked here. This, we feel, is very strong evidence that a nlethod for fine sizing could be developed to use ultrasonic waves as the separating force. Professors Sun and Mitchell indicate that floccula-ti~n will present no difficulty: "During the investigation it was found that ultrasonic waves were able to dislodge the adhering gangue particles from the min- era1 surfaces and to disintegrate the soft particles into fine sizes." Thus, it appears that the use of ultrasonic waves as the power behind a size separation process may eliminate the problem of flocculation encountered in most wet sizing techniques. It would be interesting to size a fine material under conditions in which particles are known to be flocculated and under conditions where they are known to be dispersed. Use of a sizing technique of this nature may solve several problems that have long persisted in the fine sizing field. The flocculation problem may be overcome; the continuous production of closely sized fractions may be possible with a device similar to the one described; and with photoelectric methods, rapid analyses in the subsieve sizes may be possible. In general, we suggest that this paper be followed by a study of the sizing applications of the apparatus described. We feel that such a study would not only be rewarding in providing a new sizing technique, but also that the present paper almost requires such work to complete the picture of particle motion in a field of ultrasonic waves.

Jan 1, 1958

By Bernardo Martins Braga

O curto-circuito é um dos componentes dos modelos de compartimentos utilizados na análise das curvas de Distribuição de Tempos de Residência (DTR) de distribuidores para lingotamento contínuo. Foram identificados na literatura, no entanto, diversos exemplos de seu emprego incorreto. O presente trabalho discute a interpretação física do curto-circuito, as formas de sua quantificação e as consequências nocivas de sua presença sobre a limpidez do aço e o controle da temperatura do aço no molde. Mostra-se que o método de cálculo dos parâmetros dos modelos de compartimentos deve ser alterado caso haja curto-circuito no distribuidor. É constatado que o curto-circuito deve ser evitado durante o projeto de distribuidores destinados à fabricação de aços limpos devido ao risco de perdas metálicas por macroinclusões.

Aug 16, 2017

By Edward A. Metz, R. Larry Grayson, Douglas F. Scott

We describe inconsistencies in disease and illness reporting in U.S. mining, identify under-reporting of disease and illness in U.S. mining, and summarize selected disease and illness in U.S. mining from 1983 through 2001. Methods: We summarized information on mining-related disease and illness data for the years 1983-2001 from the Mining Safety and Health Administration database (MSHA). Results: Discrepancies exist in types of information collected by the Centers for Disease and Control, the National Institute for Occupational Safety and Health, and the Mining Safety and Health Admin¬istration database. Several factors, including a worker's fear of losing his or her job, health insurance, or other job-related benefits contribute to under-reporting of disease and illness information in the US mining industry. Conclusions: Since 1997, both number of workers employed in mining and disease and illness rates have decreased; however, the highest disease and illness rates in mining continue to be coal worker's pneumoconiosis and hearing loss.

By Edinson Aponte

En la mina Yauliyacu (Huarochirí, Lima), de una profundidad aproximada de 1500 metros, existe un problema de sismicidad inducida por el minado, que puede derivar en desprendimientos o estallidos de rocas que afecten directamente a los trabajadores, equipos y a la infraestructura de la mina. Gracias al sistema microsísmico de monitoreo en tiempo real con el que cuenta esta unidad, se realizó un análisis del registro de estos eventos durante el segundo semestre del 2017 para identificar los niveles afectados o de mayor susceptibilidad a la manifestación de estallidos de rocas. A partir de ello, con el software Seis VisLite, se pudo hacer la zonificación microsísmica para la veta M. Los parámetros geomecánicos del macizo rocoso, con predominancia de arenisca silicificada, se introdujeron en el software Phase2 v.5.0 y se analizó el comportamiento de la redistribución de los esfuerzos principales en aberturas subterráneas en forma de herradura de la mina Yauliyacu, con otro tipo de abertura propuesta en la presente investigación.

Sep 20, 2019

By Fernando Machuca Pérez

"La minería genera una gran cantidad de desechos de material como lastre, ripios y escorias, que son depositados en un área sin considerar el impacto ambiental que estos producen. En ese sentido, el diseño de escombreras no considera las condiciones ambientales en la industria chilena, y por ello se privilegió el costo mínimo en depositar este lastre sin un diseño que considere esta variable. Por tal motivo, el objetivo general del texto es diseñar los algoritmos de secuencia de llenado de escombreras, considerando las siguientes variables forma, altura, talud y orientación, condiciones meteorológicas y económicas en su diseño, de tal manera que se produzca el menor impacto ambiental por la emisión de material particulado. "

Sep 1, 2015

By Edwin Anatolio Calderón Vicuña

En la actividad minera, los costos de operación en perforación y voladura son importantes puesto que el logro óptimo de una voladura óptima, no debe afectar la eficiencia física y económica de las demás operaciones (perforación, carguío, transporte, trituración). En este sentido, el presente texto nuevos diseños de perforación y voladura en las vetas angostas donde se elimine la sobreexcavación y el costo de sostenimiento.

Nov 24, 1993

By Monge G. Enrique O.

"El presente trabajo describe el proceso, elementos y las pruebas correspondientes para el diseño de plantas de tratamiento del carbón en el Perú. La construcción del diseño de plantas de tratamiento de carbón contiene cuatro soportes: las pruebas de laboratorio, especificaciones generales, planos topográficos e información económica, las cuales legitimizan su avance estructural y mejoran el producto minero. Así también su construcción requiere de cierto conocimiento íntimo del material, de las maquinarias y los procesos a emplearse. Un plano de diseño parece un acertijo en las que las unidades se aglomeran y los transportadores, elevadores, chutes, tuberías y cables forman una imagen extraña y un tanto artística. La función de este diseño de tratamiento busca contener el organismo, llenar la función y permitir la vigilancia, reparación y mantenimiento de las unidades, evitando cualquier exceso de espacio pues sería perjudicial y antieconómico. "

Jul 20, 1956

By L. S. Gormely

El presente texto describe el desarrollo del proceso de lixiviación adecuado en canchas para un yacimiento en óxidos de cobre. La lixiviación en canchas es una alternativa de tratamiento útil que se considera para aquellos óxidos cuya ley sea demasiado baja para procesarlos por trituración fina y lixiviación por agitación y que presenten dificultades técnicas para someterlos a lixiviación en tinas. Así, en el sistema de canchas, la lixiviación será más rápida y más eficiente que en una simple lixiviación en pilas.

Jan 1, 1980

By M. G. Sedlow

La mina de Udokan está situada en Siberia Oriental, distrito de Kalar de la región Chita y es uno de los mayores yacimientos minerales de cobre, donde dadas las condiciones hidrogeológicas del yacimiento se eligió el proceso de explotación a cielo abierto. En ese sentido, el presente texto señala el diseño de explotación a cielo abierto, -en tanto los cálculos mostraron que la mayor parte de las reservas pueden beneficiarse por medio de este tipo de explotación. Asimismo, sus límites se determinaron por el factor límite de estériles. Para averiguar la firmeza de los taludes de la explotación a cielo abierto, una agencia competente hizo pruebas con muestras de la roca y del mineral, así como para verificar el ángulo de inclinación más apropiado de los taludes.

May 1, 1976

By Manuel Vicente Vásquez

El presente texto describe dos posibles alternativas para el diseño de ventilación de la mina Mallay, ubicada en la provincia de Oyón, Lima. La infraestructura de la mina tiene 6 niveles, donde se ha previsto contar con 3 chimeneas principales de ventilación, cuyo requerimiento de aire mediano y largo es de 70,707 y 154,652 cfm, respectivamente. En ese sentido, se plantean dos alternativas de solución para este sistema de ventilación: La primera alternativa consiste en la construcción de dos chimeneas a superficie para la extracción del aire viciado y en el que cada chimenea contará con un ventilador principal como extractor en superficie. Mientras que la segunda alternativa, utilizará el nivel 4470 para la extracción a superficie del aire viciado mediante un ventilador principal instalado en la bocamina. Para ello, se realizará una ampliación a una sección de 10’x10’ en los primeros 70 m de la bocamina y se construirá un by pass para permitir el acceso al nivel. De las dos alternativas planteadas, ambas satisfacen el requerimiento de aire a mediano y largo plazo de la mina, con la diferencia que la segunda alternativa cuenta con la ventaja que el Nv. 4470 ya existe y sólo requiere ampliar la sección; mientras que en la primera alternativa las chimeneas tienen que ser construidas.

Sep 12, 2011

By José Adachi

La ventilación de minas es un proceso relacionado a la dinámica de la mina que genera el acondicionamiento del aire circulante a través de las labores subterráneas. Como tal, los requerimientos de la mina se adecúan de acuerdo a la estructura física de la misma. En consecuencia, no existe un sistema de ventilación estático puesto que los sistemas requieren de constantes monitoreos y ajustes para alcanzar una ventilación óptima. Así, el presente trabajo señala los aspectos importantes del diseño de ventilación principal de la mina Condestable realizado entre el período de 1992 y 1993 por encargo de la compañía minera. La metodología empleada comprende mediciones de campo y el cálculo de requerimientos del caudal. Asi también está la construcción del modelo de red de ventilación y el diseño del nuevo ciricuito con ventilación forzada.

Nov 24, 1993