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By M. A. Westbrook, J. F. Redmond

A method is presented for obtaining porosities of consolidated formations from the drill returns. The method provides a means of determining the bulk volume of a large number of particles, such as drill cuttings, by employing a capillary diaphragm for removal of surplus surface liquid from saturated cuttings having pores of capillary size, which reduces tile amount of error present in former attempts to determine the porosity of drill cuttings. The method makes it possible to obtain a considerable amount of supplementary data on consolidated formations. Introduction Porosity is a direct measure Of the volume of the voids in a formation. It may be used for estimating the permeability of a formation provided that a correlation of these two properties has been obtained from the analysis of cores from the particular formation, and it forms a desirable aid in interpreting electric logs.1 These useful applications have emphasized the importance of developing a method for determining the porosity of the formations penetrated by a well by some means other than coring, for even in exploration wells it is neither economical nor practical to core all formations. Also, in wells where at the outset the coring program appears comprehensive, it is frequently found that, after certain layers are drilled. more complete information is required for evaluation. Porosity values, if obtainable, would be of great assistance in such cases. In soft formations it is usually possible to obtain porosities after drilling with cores taken by side-wall coring devices. In hard consolidated formations, that coring procedure is not alway practicable and other methods for determining porosities must be resorted to. A first choice is the examination of drill cuttings. if an analytical procedure is available that is both accurate and reasonably rapid. Such a procedure has been devised and is applicable to sand and limestone cuttings having pores of capillary size. Although imp,ovements can yet be made, it is believed that the new technique is sufficiently well developed to warrant description, Preparation of Sample The well samples containing the cuttings representative of the particular intervals are first washed through a No' 6 gauge screen. The cuttings remaining on the Screen are large enough to use lor the porosity measurements' From the sized sample, the cuttings to be examined are separated from the shale and Other extraneous particles. It is recommended that 10 to 15 c.c. of the selected cuttings be used for measurements, although only 4 to 5 c c. were used to make the comparative tests described later in this article. These cuttings are extracted and dried by the same methods used for preparing core samples for porosity and other measurements. The separation, which is at present being carried out by hand, is the most tedious

Jan 1, 1946

By Henry Fay

LAWRENCE ADDICKS, New York, N. Y., (communication to the Secretary *).-There are two points regarding gun erosion on which I want to say a few words: The first is about the analogy to hardening of wire by drawing and the second the possible role played by thermal conductivity of the metal. The ductile metals do not necessarily become brittle when cold-worked as it is possible to draw copper from 3. in. in diameter soft rod down to the finest hair wire without annealing. This results, as stated in the paper, informing a hardened core with a very hard skin. The electrical conductivity is depressed some 3 per cent. and the tensile strength nearly tripled. A soft copper wire 0.08 in. in diameter and 6 in. long may be twisted 150 or more turns before the wire will snap under torsion. A hard-drawn wire will break after perhaps a dozen twists, the skin breaking down at one spot. Just wiping the wire with a cloth clamp with dilute nitric acid has a remarkable effect in increasing the number of twists without appreciably lowering the tensile strength, showing the skin to be very thin. On the other hand, pickling the rod, from which the wire is to be drawn, in dilute nitric acid also greatly increases the number of twists which can be applied to the resulting wire. This indicates that the copper oxide scale produced when the wire bar is rolled hot is more or less rolled into the rod and may have much to do with the skin effect. We also know that if the conditions of drawing are not correct, the heat of deformation in the die may produce quite an annealing effect. The application of this analogy to what occurs in a gun to produce the effects noted offers some difficulty. The second point is that as the temperature of the surface of the gun bore has undoubtedly a great deal to do with the amount of erosion, one natural remedy would be to conduct the heat away fast enough to keep this temperature within desired limits. This brings up the question of the thermal conductivity of the metal of which the gun is made. We know that the conductivity of pure metals is greatly altered by alloying and that the alloy is usually, though not always, lower in conductivity than the metal before alloying. This is interesting when it is considered that the high-nickel and manganese-steel alloys show greater erosion as stated. Would it not be instructive to study the question from the point of view of heat conducted away from the eroded surface?

Jan 3, 1917

By A. W. Fahrenwald

Cell depth has been for many years a controversial question in a flotation-machine performance. In the impeller type of machine, we are really talking about impeller submergence—i.e., the depth in the pulp at which the impeller is required to produce aeration. The Flotation Machine a Compressor A flotation machine can be considered as an air compressor or blower in which the aeration mechanism takes in air under atmospheric pressure† (14.7 Ib. per sq. in.) and blows it into the pulp (the pulp being the receiver) against a head of 14.7 plus cell depth in feet times 0.4333 lb. per sq. inch. In commercial machines the impeller is submerged in the pulp to depths ranging from about 2 ft. in the smaller machines to as much as 5 ft. in some of the larger ones. In the 2-ft. cell, the hydrostatic head on the impeller is (2 X 1.25 =) 2.50 ft. (assuming a pulp density of 1.25), equal to (2.50 X 0.4333 =) 1.078 Ib. per sq. in. For the 5-ft. cell, the hydrostatic head against which the impeller must make air is (5 X 1.25 X 0.4333 =) 2.708 Ib. per sq. inch. In the small 2-ft. machine, the volume reduction is (1.078 ÷ 0.147 + 0.01078 =) 6.85 per cent and in the large, 5-ft. machine, the volume reduction is (2.708 0.147 + 0.02708 =) 15.5 per cent. It is obvious that for unit volume of free air intake the larger machine receives a lesser (8.65 per cent less) volume of aeration* than the smaller machine. The bubbles in escaping from the pulp expand and are greater in volume at the surface than at the moment of their formation in the pulp; 7.33 and 18.4 per cent larger, respectively, in the small and large machines. Just how this situation affects flotation is not known. It is interesting to note, however, that the bubbles are smaller in the zone of the coarsest particles and largest in the upper zones of the her, more easily suspended mineral particles. It would seem that the relationship might be reversed advantageously. The deep cell is at a greater disadvantage in this respect than the shallow cell. Nothing, however, presumably can be done about this situation. The shallow cell is a partial corrective. Not only does the compression ratio of aeration increase with impeller submergence but the hydrostatic head against which the impeller must operate also increases. The Flotation Machine as a Pump Since, in impeller machines that rely upon their own operation to produce aeration, pulp Passes through the im-

Jan 1, 1947

By C. M. Romanowitz, H. A. Sawin

PLACER dredging as we know it today, especially gold dredging, is an industry about 40 years old, dating from the beginning of this century, when a few mining men in California saw the possibilities in adapting earlier dredging methods to California placer fields in and immediately adjacent to streams. At first it was thought that dredges must be operated in existing stream beds, but experience soon pointed the way to dredging alluvial deposits of old stream beds, operating on landlocked dredging ponds. Within the first five or six years of modern dredging, experience was gained so rapidly that dredges to dig 60 ft. below water level were designed and built. This depth was believed by many operators to be the maximum at which successful dredging could be carried on. Alluvial deposits of gold, suitable for dredging, were readily handled by equipment available and developed early in the industry's history. Much of the gravel was easily dug and if deposits were cemented, contained large boulders, or for some other reason could not be handled by a dredge, the ground was abandoned in favor of other deposits that could be handled easily. Many failures, adversely affecting the dredging industry, resulted from attempts at that time to extend the scope of placer dredging with equipment not suited to the formation being dredged. Dredging ground suited to early equipment was exhausted eventually, but many supposedly undredgeable alluvial fields containing gold in commercial quantities were found and became known to dredgemen. Improved equipment was needed to work deposits known to exist at great depths or associated with unusual obstructions. Operating companies looked to dredge manufacturers and metallurgists, particularly manganese-steel foundries, for improved designs and materials. Experience developed in California has spread throughout the world, and dredges of improved types have been adapted to the mining of platinum and tin as well as gold. Space will not permit a description of all the improvements in dredging since the early days, but a few of the recent developments are worthy of note, because they make possible the dredging of placer deposits of great depth, those having hard, irregular bedrock or containing large boulders in great quantities, and compacted or cemented gravel deposits. DEEP DREDGING At Hammonton, Calif., Yuba Consolidated Gold Fields has, within the past two years, built its dredge No. 20. The company has operated continuously in that field since 1904, digging successively at maximum depths of 60 ft., 8o ft., 112 ft., and now 124 ft. below water level. The first dredges were equipped with buckets of 6 and 7 ½ -cu- ft. capacity; six dredges operating in the field today have 18-cu. ft. buckets, digging now 65 to 124 ft. below water level. Some ground near Hammonton has been dug for the third time, owing to the devel-

Jan 1, 1941

HeRbeRT Haas,—(communication to the Secretary*).—Fig. I plainly shows that the comparison of the steam-turbine and Diesel-engine plants was made on a basis of 6,000 kw. continuous operating load. The tabulation of costs for an 8,000-kw. load is merely done to show the relation of increased output on the unit power costs. Both plants would sacrifice their reserve capacity, which is to insure continuous operation, if operated to generate 8,000 kw. continuously; unless Mr. Hawkins advocates the operation of one of the turbines at 33.33 per cent. overload, retaining the other as a standby, which, besides overtaxing the turbine, results in a great loss of efficiency. It would be well for Mr. Hawkins to qualify certain of his statements. Thus, under the caption "Diesel Engines, Selection of Units," he says that '(two of the largest Diesel-engine builders take the position that, in the present state of the art, they would not attempt to put out units over 1,000 kw.;" he should qualify it by saying that "two of the largest American Diesel-engine builders take the position that, in the present state of the art in the United States, they would not attempt to put out units over 1,000 kw." The reason that American manufacturers restrict themselves at present to units of that size is to be found not so much in a lack of experience in building large units as in a lack of a sufficient demand for them in the United States to justify the heavy capital outlay connected with the development of such engines. There are relatively few localities in the United States where fuel prices are as high as they are in the South-, west, and regions of high fuel prices are sparsely settled and have a very limited demand for power. Where low fuel prices obtain, steam turbines are active competitors of Diesel engines for large power plants, particularly when the station load factor is low. As most of the American manufacturers of Diesel engines work under license agreements with European manufacturers, which limit their market to the United States and its possessions and Canada, a further limitation is imposed upon them, leaving to European firms practically the entire world as a market. The greater number of power plants range in size from 300 to 1,000 hp.; the average size of the central station, on a basis of all central stations in the United States, not greatly exceeding 500 kw.; in these sizes it is hard for steam engines to compete with Diesel engines, unless fuol oil prices are excessively high, and coal prices very low, or the exhaust steam is used for heating purposes.

Jan 1, 1917

By A. W. Fahrenwald

Cell depth has been for many years a controversial question in a flotation-machine performance. In the impeller type of machine, we are really talking about impeller submergence—i.e., the depth in the pulp at which the impeller is required to produce aeration. The Flotation Machine a Compressor A flotation machine can be considered as an air compressor or blower in which the aeration mechanism takes in air under atmospheric pressure† (14.7 Ib. per sq. in.) and blows it into the pulp (the pulp being the receiver) against a head of 14.7 plus cell depth in feet times 0.4333 lb. per sq. inch. In commercial machines the impeller is submerged in the pulp to depths ranging from about 2 ft. in the smaller machines to as much as 5 ft. in some of the larger ones. In the 2-ft. cell, the hydrostatic head on the impeller is (2 X 1.25 =) 2.50 ft. (assuming a pulp density of 1.25), equal to (2.50 X 0.4333 =) 1.078 Ib. per sq. in. For the 5-ft. cell, the hydrostatic head against which the impeller must make air is (5 X 1.25 X 0.4333 =) 2.708 Ib. per sq. inch. In the small 2-ft. machine, the volume reduction is (1.078 ÷ 0.147 + 0.01078 =) 6.85 per cent and in the large, 5-ft. machine, the volume reduction is (2.708 0.147 + 0.02708 =) 15.5 per cent. It is obvious that for unit volume of free air intake the larger machine receives a lesser (8.65 per cent less) volume of aeration* than the smaller machine. The bubbles in escaping from the pulp expand and are greater in volume at the surface than at the moment of their formation in the pulp; 7.33 and 18.4 per cent larger, respectively, in the small and large machines. Just how this situation affects flotation is not known. It is interesting to note, however, that the bubbles are smaller in the zone of the coarsest particles and largest in the upper zones of the her, more easily suspended mineral particles. It would seem that the relationship might be reversed advantageously. The deep cell is at a greater disadvantage in this respect than the shallow cell. Nothing, however, presumably can be done about this situation. The shallow cell is a partial corrective. Not only does the compression ratio of aeration increase with impeller submergence but the hydrostatic head against which the impeller must operate also increases. The Flotation Machine as a Pump Since, in impeller machines that rely upon their own operation to produce aeration, pulp Passes through the im-

Jan 1, 1947

By Bunting S. Crocker

DISCUSSION Fred C. Bond (Allis-Chalmers Manufacturing Co., Milwaukee)—This is a very complete description of the screened ore fine grinding installation at Lake Shore. The only important data which might be desired are the operating screen analyses and power consumption in each stage. The statement that the energy consumption, or kw-hr per ton, remains constant for all mill diameters is of great interest, but the opinion is widely held that the kw-hr per ton decreases as the mill diameter increases. More definite test work is needed to establish this point. Screened ore fine minding has been successful on hard Lake Shore ore &d on about equally hard South African Rand ore. However, it is still not determined whether it would be successful on ores of average or less than average hardness. It is possible that with softer ores the consumption of ore pebbles might be exceedingly high, or might even exceed the available supply. Also, the softer feed might result in a rapid buildup of harmful intermediate sizes in the mill circuit. These factors can be determined only by a full scale or long pilot mill test. In any case the Lake Shore staff is to be congratulated for this pioneering work. Bunting. S. Crocker (author's reply)—As requested, the operating screen analyses and Bower consumption for each of the four stages of pebble milling are given, see Table XII. The fist three stages are in series circuit; the last stage is in closed circuit. The primary circuit had rock feed averaging 124 g per piece and the final stages rock feed weighing 63 g. The intermediate stages had mixtures of the two. Regarding Mr. Bond's statement that "the opinion is widely held that the kw-hr per ton decreases as the mill diameter increases," I would say that I have frequently been told this but have never yet been shown any evidence or seen any evidence published that would prove this statement. It may be true that some of the big mills recently installed have mechanical power savers such as micarta bearings and better lubrication, but as far as the power consumed inside the mill is concerned, we at Lake Shore think that it varies as D2 .", and that is all. We published extensive information on this point in our 1939 paper. Recently Mr. Banks at Consolidated published the figures on his 10 ft 8 in. x 12 ft 0 in. rod mill. As reported in this discussion by the writer: we calculated the power necessary to run this mill based on our 7x6-ft rod mill to within 1 pct of the actual power reported by Mr.

Jan 1, 1953

The Carol concentrator of the Iron Ore Company of Canada was completed in 1962 and through expansions .has become one of the largest iron ore beneficiation plants in the world with a new feed capacity of 147,000 It (149,350 mt) per day. The Carol ore contains specular hematite and magnetite having a bulk density of 166 lb/cu ft (2660 kg/m3) It is crushed to 7 in. (178 mm) for feeding to autogenous mills for subsequent concentration by Humphrey’s s spirals. Mine ore is loaded into 150 It (152 mt) trucks by 10 cu yd (7.65 m3) shovels and then, at a transfer station, is loaded into side-dump rail cars of 100 It (101.6 mt) capacity. Each train is composed of twenty cars and operates completely automatically as it hauls the ore to the primary crushing station, dumps its loads, and returns to the loading station, a 11.5 mile (18.5 km) round trip. The primary crushing station is composed of two 60 in. by 89 in. gyratories installed in 1962 and 1972. Neither crusher has a grizzly ahead of it. Rail cars are dumped into 300 It (305 mt) pockets which contain splitters for spider protection. What few oversize boulders which do not go through the crushers are drilled and blasted. Liners in the dump pocket and the crushed ore pokcet under the crusher are manganese steel. Both crushers are direct driven by 700 hp (5 15 kw) motors, and the hardness of the ore causes them to pull almost their full power. Crushed ore drops into a 160 It (163 mt) pocket equipped with radioactive level control units for low-intermediate, and high level indications. The low level unit stops and starts the apron feeder, the intermediate unit controls the feeder speed, and a high-level reading stops the car dumping. Apron feeders are 72 in. (1. 82 m) wide and are variable speed from 12 to 52 fpm (3. 7 to 15. 8 mpm). They discharge to 60 in. (1.52 m) wide takeaway belts. Crushing capacity, nominal and maximum for each crusher, is 4, 000 ltph (4. 064 mtph) and 4,600 ltph (4,674 mtph). The crusher eccentric is removed by a conventional cart which travels on removable rails. Mainshaft removal and maintenance is done by 65 and 75/10 ton cranes. Mainshaft replacement requires eight hours, a usual time, but concave replacement requires only sixteen hours, a result of a fine maintenance organization. Mantle relining is done at grade next to a repair track. Zinc is used for backing both mantles and concaves. Both the electrical rooms and the dust collectors are located below grade,

Jan 1, 1978

By Bunting S. Crocker

DISCUSSION Fred C. Bond (Allis-Chalmers Manufacturing Co., Milwaukee)—This is a very complete description of the screened ore fine grinding installation at Lake Shore. The only important data which might be desired are the operating screen analyses and power consumption in each stage. The statement that the energy consumption, or kw-hr per ton, remains constant for all mill diameters is of great interest, but the opinion is widely held that the kw-hr per ton decreases as the mill diameter increases. More definite test work is needed to establish this point. Screened ore fine minding has been successful on hard Lake Shore ore &d on about equally hard South African Rand ore. However, it is still not determined whether it would be successful on ores of average or less than average hardness. It is possible that with softer ores the consumption of ore pebbles might be exceedingly high, or might even exceed the available supply. Also, the softer feed might result in a rapid buildup of harmful intermediate sizes in the mill circuit. These factors can be determined only by a full scale or long pilot mill test. In any case the Lake Shore staff is to be congratulated for this pioneering work. Bunting. S. Crocker (author's reply)—As requested, the operating screen analyses and Bower consumption for each of the four stages of pebble milling are given, see Table XII. The fist three stages are in series circuit; the last stage is in closed circuit. The primary circuit had rock feed averaging 124 g per piece and the final stages rock feed weighing 63 g. The intermediate stages had mixtures of the two. Regarding Mr. Bond's statement that "the opinion is widely held that the kw-hr per ton decreases as the mill diameter increases," I would say that I have frequently been told this but have never yet been shown any evidence or seen any evidence published that would prove this statement. It may be true that some of the big mills recently installed have mechanical power savers such as micarta bearings and better lubrication, but as far as the power consumed inside the mill is concerned, we at Lake Shore think that it varies as D2 .", and that is all. We published extensive information on this point in our 1939 paper. Recently Mr. Banks at Consolidated published the figures on his 10 ft 8 in. x 12 ft 0 in. rod mill. As reported in this discussion by the writer: we calculated the power necessary to run this mill based on our 7x6-ft rod mill to within 1 pct of the actual power reported by Mr.

Jan 1, 1953

By H. Q. Golder

With the development of open-pit mines of ever increasing depth, the stability of the pit slopes in soil and rock is of growing economic importance. How is it, considering the newness of the problem that we know anything at all about the stability of man-made slopes? How much do we know and how did it all begin? We can make a reasonable estimate of the stability of a slope in soil, but we know as yet woefully little about slopes in rock because the problem is much more complex. It is only with the development of the capacity of the computer to handle wearisome calculations that any but the simplest of stability problems in rock have become possible of solution. But what we do know is a guide of immense value in the design of slopes in rock. But remember that the computer is a great elevator; it takes everyone's feet off the ground! So our study began with soils and what is now known as "Soil Mechanics", which is a translation of the German term "Bodenmechanik". It is a pity that the title of Terzaghi's book "Erdbaumechanik" was not used instead. But here the task of the translator is difficult because English lacks the facility of German to string together several words and make a new one whose meaning is unequivocal. Thus, we have the three German words 'Erde' 'Bau' and 'Mechanik'. 'Erde' means "earth" in almost all the English senses including "soil". 'Bau' and 'bauen' convey the meaning of building and construction, and 'Mechanik' means "mechanics" which includes statics and dynamics and the 'mechanics of materials', etc. So that a correct translation of the German is a phrase like "the mechanics of construction with, in or on, earth or soil." However we are stuck with 'Soil Mechanics' and that is where it all began. Now this study or subject did not begin with 'Terzaghi' although his contribution to the subject is unquestionably greater than that of any other single person. Every scientific discipline has its 'classical problem' and that of soil mechanics is 'the earth pressure problem' and it began in the 17th century. In fact, it probably began before this with the introduction of gunpowder to Western Europe in about the 15th century, but the first known reference to the problem was by Vauban, a distinguished French military engineer who lived from 1633 to 1707 and who gave rules for the dimensions of retaining walls from 6 to 80 ft. in height. The soil properties were not mentioned by Vauban probably because the rules were intended to be applied to fortifications only, but Poncelet (1840) claimed that Vauban's rules were based on a theory, and not simply on military and practical

Jan 1, 1972

By H. A. Sawin, C.M. Romanowitz

Placer dredging as we know it today, especially gold dredging, is an industry about 40 years old, dating from the beginning of this century, when a few mining men in California saw the possibilities in adapting earlier dredging methods to California placer fields in and immediately adjacent to streams. At first it was thought that dredges must be operated in existing stream beds, but experience soon pointed the way to dredging alluvial deposits of old stream beds, operating on landlocked dredging ponds. Within the first five or six years of modern dredging, experience was gained so rapidly that dredges to dig 60 ft. below water level were *designed and built. This depth was believed by many operators to be the maximum at which successful dredging could be carried on. Alluvial deposits of gold, suitable for dredging, were readily handled by equipment available and developed early in the industry's history. Much of the gravel was easily dug and if deposits were cemented, contained large boulders, or for some other reason could not be handled by a dredge, the ground was abandoned in favor of other deposits that could be handled easily. Many failures, adversely affecting the dredging industry, resulted from attempts at that time to extend the scope of placer dredging with equipment not suited to the formation being dredged. Dredging ground suited to early equipment was exhausted eventually, but many supposedly undredgeable alluvial fields containing gold in commercial quantities were found and became known to dredgemen. Improved equipment was needed to work deposits known to exist at great depths or associated with unusual obstructions. Operating companies looked to dredge manufacturers and metallurgists, particularly manganese-steel foundries, for improved designs and materials. Experience developed in California has spread throughout the world, and dredges of improved types have been adapted to the mining of platinum and tin as well as gold. Space will not permit a description of all the improvements in dredging since the early days, but a few of the recent developments are worthy of note, because they make possible the dredging of placer deposits of great depth, those having hard, irregular bedrock or containing large boulders in great quantities, and compacted or cemented gravel deposits. Deep Dredging At Hammonton, calif., yuba consolidated Gold Fields has, within= the past two years, built its dredge No. 20. The company has operated continuously in that field since 1904, digging successively at maximum depths of 60 ft., 80 ft., 112 ft., and now 124 ft. below water level. The first dredges were equipped with buckets of 6 and 7½-cu. ft. capacity; six dredges operating in the field today have 18-cu. ft. buckets, digging now 65 to 124 ft. below water level. Some ground near Hammonton has been dug for the third time, owing to the devel-

Jan 1, 1943

By N. N. Breyer, W. R. Warke, S. Dinda

A model for the low-temperature peak, LTP, in fcc metals is presented. In high stacking fault energy metals, e.g., aluminum and nickel, it is proposed that a reversible dislocation glide and climb process might give rise to the observed high relaxation strengths and activation energies equivalent to volume self-diffusion values. In low stacking fault energy metals such as copper, gold, and silver, dislocation climb is energetically unfavorable at T < 0.5 Tm and the relaxation process is therefore mainly one of reversible glide. This model is used to explain the effect of a change in grain size and impurities on the LTP. THE low-temperature grain boundary damping peak, LTP, in spectroscopically pure aluminum,&apos; copper,17&apos; gold,2 nickel,3 and silver4 has recently been the subject of a detailed investigation by our laboratory. The main results are summarized in Table I. An empirical relationship was found between the peak relaxation strength, AE, and the width of the dissociated dislocation, do,* given by: do = µa2/24p? (Ref. 5) [I] where p is the shear modulus, y is the stacking fault energy,6-8 and a is the lattice parameter, Table 11. 4E is plotted against do in Fig. 1. Two regions are apparent: region (a)! which contains all but one of the metals, where AE decreases linearly with increasing do, i.e,, AE 1/do; and region (b), where ?E is approximately independent of do. A similar trend is observed when the ratio Qp/QL is plotted as a function of do, Fig. 2. (Qp and Ql are the activation energies for the peak and for lattice diffusion, respectively). In the past, when Qp and QL were found to be comparable, the relaxation peak was explained in terms of grain boundary sliding.10 The equivalence of Q, and Qb. the activation energy for grain boundary diffusion, prompted the suggestion that it was a consequence of reversible grain boundary migration on an atomic scale." A new model for the LTP, which is based on the experimental relationships in Figs. 1 and 2, is presented. In region (a) it is suggested that stress relaxation occurs through the reversible glide and climb of grain boundary dislocations; in region (b) only reversible glide can take place. The resultant effect on the grain boundary might be sliding or "a bowing out", but, since the total distance moved will be 50A, the atomic movements are considered of primary im- portance. It must be emphasized that this is only a qualitative model, but at the moment it appears the most likely way to explain the widely varying, but re-

Jan 1, 1969

By C. M. Romanowitz, H. A. Sawin

Placer dredging as we know it today, especially gold dredging, is an industry about 40 years old, dating from the beginning of this century, when a few mining men in California saw the possibilities in adapting earlier dredging methods to California placer fields in and immediately adjacent to streams. At first it was thought that dredges must be operated in existing stream beds, but experience soon pointed the way to dredging alluvial deposits of old stream beds, operating on landlocked dredging ponds. Within the first five or six years of modern dredging, experience was gained so rapidly that dredges to dig 60 ft. below water level were *designed and built. This depth was believed by many operators to be the maximum at which successful dredging could be carried on. Alluvial deposits of gold, suitable for dredging, were readily handled by equipment available and developed early in the industry&apos;s history. Much of the gravel was easily dug and if deposits were cemented, contained large boulders, or for some other reason could not be handled by a dredge, the ground was abandoned in favor of other deposits that could be handled easily. Many failures, adversely affecting the dredging industry, resulted from attempts at that time to extend the scope of placer dredging with equipment not suited to the formation being dredged. Dredging ground suited to early equipment was exhausted eventually, but many supposedly undredgeable alluvial fields containing gold in commercial quantities were found and became known to dredgemen. Improved equipment was needed to work deposits known to exist at great depths or associated with unusual obstructions. Operating companies looked to dredge manufacturers and metallurgists, particularly manganese-steel foundries, for improved designs and materials. Experience developed in California has spread throughout the world, and dredges of improved types have been adapted to the mining of platinum and tin as well as gold. Space will not permit a description of all the improvements in dredging since the early days, but a few of the recent developments are worthy of note, because they make possible the dredging of placer deposits of great depth, those having hard, irregular bedrock or containing large boulders in great quantities, and compacted or cemented gravel deposits. Deep Dredging At Hammonton, calif., yuba consolidated Gold Fields has, within= the past two years, built its dredge No. 20. The company has operated continuously in that field since 1904, digging successively at maximum depths of 60 ft., 80 ft., 112 ft., and now 124 ft. below water level. The first dredges were equipped with buckets of 6 and 7½-cu. ft. capacity; six dredges operating in the field today have 18-cu. ft. buckets, digging now 65 to 124 ft. below water level. Some ground near Hammonton has been dug for the third time, owing to the devel-

Jan 1, 1943

By J. L. Wyatt

The electrical resistance of titanium as a function of purity and temperature was measured from —325" to 2800°F. Two points of inflection in the data plots were found, and an increase in resistance with increase in temperature above the transformation was observed. IN 1948 Greiner and Ellis&apos; reported on the thermal and electrical properties of titanium, including a study of the effect of temperature on the electrical resistance. The materials used in the research were obtained from the United States Bureau of Mines, where powder metallurgy techniques were being employed for consolidation purposes. These authors found a decreasing resistance with increasing temperature above 1625°F, and concluded that the latter temperature represented the point of transformation from hexagonal close-packed to body-centered cubic structure on heating. They reported a specific resistivity value of 55x10-6 ohm-cm, and showed that dp/dt was positive, decreasing with increasing temperature above ambient. In one of his early texts Hume-Rothery2 mentioned that certain metals, including titanium, exhibited minima in plots of resistance vs temperature. He indicated that this occurred at 212°F. Reporting many years ago on the effect of temperature on the electrical resistance of zirconium, de Boer and Fast³ showed that phase transformation in this sister metal occurred over a range of temperatures, the extent varying with its purity. Graphs of resistance vs temperature, showing transformation effects over a range of several hundred degrees, were presented for relatively impure metal. In the latter case the resistance decreased with increasing temperature throughout the region of partial phase transformation, thereafter increasing as predicted for normal metals by modern theories of the solid state. Since the quality of metal used in previous studies was substandard compared with iodide-type titanium, and in view of the fact that there seemed to be some rather unusual observations in the reported literature, it was decided to undertake a more extensive study of the thermal effects using high purity titanium, and to include temperature ranges beyond those which had been studied in order to ascertain whether or not the characteristics of zirconium would be duplicated. Experimental Materials Materials used in this investigation included commercial quality titanium (Ti-75A produced by the Titanium Metals Corp. of America) in the form of ¼ in. diam rods, and high purity iodide titanium (produced by the Foote Mineral Co.) in the form of crystal bar of irregular size. The latter was cold-swaged into rod form and machined to the desired size. Lead wires to all specimens consisted of 18 gage Ti-75A. Chemical analyses of the stock materials are given in Table I. Experimental Procedures Specimens 2½ to 3 in. in length were prepared with diameters nominally 0.15 to 0.20 in. Two holes,

Jan 1, 1954

The Molycorp operation at Mountain Pass, California, is unique in that it is the only place in the world where bastnasite is mined. The orebody is a metamorphic carbonatite; the mineral of importance is bastnasite, a fluocarbonate containing principally cerium, lanthanum, neodymium, and praseodymium, with small amounts of samarium, gadolinium, and europium. In addition, appreciable quantities of barite, calcite, and strontianite are present. Molycorp had mined and processed the bastnasite ore since the early 1950s at a rate of about 100 tpd by a unique hot flotation process in a plant originally used for cyanidation of gold ore. The concentrate was primarily used in high quality glass polishing, lighter flints, ceramics, and in specialty steel making. Then in the early 1960s, Molycorp and the Colorado School of Mines Research Foundation developed a solvent extraction process to recover europium from the bulk rare earth concentrates. In 1965 a new SX plant was constructed to produce this material for use as a red phosphor in color television tubes. Because of the demand for this material, a new 500 tpd concentrator was designed and constructed in 9 ½ months. Then, because of a sharply increased demand for europium, the plant was expanded to 1200 tpd only four months after completion of the original construction. The crushing section of the new plant was designed similarly to those portable units used in road construction and arrived on the site in ten major pieces. Foundations were concrete piers; and after their completion, the entire crushing plant composed of primary and secondary crushers, conveyors, and a screen was erected in fourteen days by a crew of ten men assisted by a mobile crane. Pit-run ore is loaded into 35-ton trucks by a 7 yd (5.4 cu m) loader, trans- ported to .the crushing plant, and dumped into the 65 -ton concrete pocket. A 42 in. by 20 ft (1.07 m by 6. 1 m) apron feeder feeds ore to the 30 in. by 42 in. (760 rnrn by 1070 mm) single toggle jaw crusher. Crushed ore is trans- ferred by a scissors conveyor to a 5 ft by 14 ft (1.52 m by 4. 27 m) double- deck screen which produces a final ¾ in. (19 rnrn) product with the oversize being treated in a 4 ¼ ft (1. 3 m) standard cone crusher. Cone crusher product is returned to the first scissors conveyor to join the primary crusher product. Dust collection is by a wet scrubber, and an operator&apos;s control room is located to overlook the primary. The entire crushing is in the open, and a 30-ton mobile crane is used for maintenance. Molycorp’s two-stage crushing plant is an excellent example of functional design for a small mining and milling operation.

Jan 1, 1978

By H. W. Straley

Some ten years ago the writer1 made a survey of college catalogues to determine what sort of training geologists were receiving in basic sciences. In the light of this compilation and subsequent experience, he made the statement at the 1939 meeting of the Institute that apparently only in schools offering a curriculum in geological engineering or in the geological option of mining engineering curricula could geologists hope to receive the training advocated by Hubbeet.2 In the former study, schools of engineering were excluded because the preparation was known beforehand. The present study aims to consider engineering schools in an effort to determine not only what basic studies are required but what use is made of them in geological courses of the third and fourth years. The present study combined the method of consulting catalogues with a questionnaire. Letters were mailed to officials responsible for supervision of mineral industries curricula requesting information On&apos; (I) the extent to which the engineering preparation of the first two years is actually used in the work of the last two; (2) the extent to which the engineering of the last two years is used in courses in geological engineering or geology; and (3) whether geologists or geological engineers teach descriptively or analytically. Both methods of securing information are open to criticism; the former on the basis of the incomplete information supplied by university catalogues concerning prerequisites and curricula and the latter on the ground that the questions may be phrased poorly or given insufficient consideration by replying officials. Fifty-six questionnaires were mailed and 39 (70 per cent) replies were received. Of the replies, 6 (11.3 per cent) were from colleges of liberal arts unconnected with engineering schools, 11 (20 per cent) were from schools having curricula in geological engineering and 22 (39 per cent) were from departments of mining engineering. The number of replies and their distribution were satisfactory in spite of the receipt of only six from liberal arts colleges. It should be understood that the universities of which the catalogues were examined were not necessarily the same in both compilations. No departments that are admittedly or known to be acting solely as service departments were recorded in the second study. Table I shows that all professional schools require mathematics through calculus, whereas only 7 per cent of the liberal arts curricula expect the same level of preparation; only 73 per cent require trigonometry and 40 per cent college "algebra. The proportion of geological engineering curricula requiring differential equations is higher than mining. All of the engineering schools require general and analytical chemistry, whereas only 77 per cent of liberal arts curricula require general and 18 per cent analytical. Twenty-two per cent of the mining and 40 per cent of the geological engineering

Jan 1, 1946

By H. W. Straley

Some ten years ago the writer1 made a survey of college catalogues to determine what sort of training geologists were receiving in basic sciences. In the light of this compilation and subsequent experience, he made the statement at the 1939 meeting of the Institute that apparently only in schools offering a curriculum in geological engineering or in the geological option of mining engineering curricula could geologists hope to receive the training advocated by Hubbeet.2 In the former study, schools of engineering were excluded because the preparation was known beforehand. The present study aims to consider engineering schools in an effort to determine not only what basic studies are required but what use is made of them in geological courses of the third and fourth years. The present study combined the method of consulting catalogues with a questionnaire. Letters were mailed to officials responsible for supervision of mineral industries curricula requesting information On&apos; (I) the extent to which the engineering preparation of the first two years is actually used in the work of the last two; (2) the extent to which the engineering of the last two years is used in courses in geological engineering or geology; and (3) whether geologists or geological engineers teach descriptively or analytically. Both methods of securing information are open to criticism; the former on the basis of the incomplete information supplied by university catalogues concerning prerequisites and curricula and the latter on the ground that the questions may be phrased poorly or given insufficient consideration by replying officials. Fifty-six questionnaires were mailed and 39 (70 per cent) replies were received. Of the replies, 6 (11.3 per cent) were from colleges of liberal arts unconnected with engineering schools, 11 (20 per cent) were from schools having curricula in geological engineering and 22 (39 per cent) were from departments of mining engineering. The number of replies and their distribution were satisfactory in spite of the receipt of only six from liberal arts colleges. It should be understood that the universities of which the catalogues were examined were not necessarily the same in both compilations. No departments that are admittedly or known to be acting solely as service departments were recorded in the second study. Table I shows that all professional schools require mathematics through calculus, whereas only 7 per cent of the liberal arts curricula expect the same level of preparation; only 73 per cent require trigonometry and 40 per cent college "algebra. The proportion of geological engineering curricula requiring differential equations is higher than mining. All of the engineering schools require general and analytical chemistry, whereas only 77 per cent of liberal arts curricula require general and 18 per cent analytical. Twenty-two per cent of the mining and 40 per cent of the geological engineering

Jan 1, 1946

By O. L. Culberson, J. J. McKetta

INTRODUCTION The equilibrium constants for methane and for water, and for ethane and water have been calculated from experimental data for the two binary systems.2,3,11,12 These constants are for the two-phase systems for the temperature range of 100" to 340°F and the pressure range of 200 to 10,000 psia. The constants for ethane are greater than those for methane. Equilibrium constants for a natural gas from a natural gas-water system6 are about the same as those for methane. Equilibrium constants for water in all three systems are very nearly the same over the greater part of the range of temperatures and pressures studied. In ideal solutions, the vapor-liquid equilibrium constants for a component should be the same in all such solutions at a given temperature and pressure.5 Few solutions are ideal, and consequently the use of generalized constants often leads to error in situations where they do not actually apply. In decidedly non-ideal systems, it is desirable to have equilibrium constant data for the specific system. Vaporization equilibrium constants may be calculated either by thermodynamic methods or from experimental vapor-liquid equilibrium data. In the thermodynamic calculation. the equilibrium constant K may be shown to be" K = fP,/fl&apos;..........(1) where fPi is the fugacity of the pure component i at the given temperature and at the vapor pressure of the pure component, and fp is the fugacity of pure i in the gaseous state at the temperature and total pressure of the system. At best the thermodynamic calculation results in a K applicable to ideal solutions. Equilibrium constants calculated from Equation (1) have been found to have only limited agreement with experimental values, even though the solutions for which they were calculated approached ideality."&apos;" If the more volatile component is at a temperature above its critical, the accuracy of the equilibrium constant for the component is made more doubtful by the use of the vapor pressure term under such conditions. In order to make the thermodynamic calculation on a component above its critical temperature, one must resort to one of the empirical methods of extrapolation of the vapor pressure to conditions above its critical point.13 Application of the thermodynamic equilibrium constant then to the methane-water or the ethane-water systems would be risky since the systems are non-ideal, and the methane and ethane are above their critical in most petroleum operations. ETHANE-WATER SYSTEM In the temperature range of from 100° to 340°F and for pressures up to 10,000 psia, the vapor-liquid equilibrium data &apos;References given at end of paper. Manuscript received in the office of the Petroleum Branch March 16, 1951.

Jan 1, 1951

By F. C. von der Weid

The mining district of Sao Fidelis, in the northern part of the State of Rio de Janeiro, is situated on the Paraiba River, where it crosses the Serra do Mar (Mountain of the Sea), in the center of a mountainous country having numerous graphitic outcrops. For a long time, these outcrops were irregularly explored by local people, who had little means and no technical advice. A few years ago, the Cia. Brasileira de Minerasgo de Grafite took over the control of some of the more important outcrops and entrusted to me the task of studying the ores and establishing a flowsheet for a pilot plant to be situated in São Fidelis—the ore to be concentrated being mostly of a very high grade. Mineralization and Ores The rocks of the district are mostly gneiss and crystalline schists, with occasional interstratification of dolomitic limestone belonging to the Archean Brazilian Fundamental Complex. The general direction of the beds is roughly between 60" and 80°E., with variable dip, generally nearly vertical. Mostly, direction and dip of the graphitic outcrops is concordant with the embanking gneisses or schists. These outcrops seem to be aligned in two or three parallel zones (Pádua, São Fidelis, Mac- abu), the Sao Fidelis zone being the best known and the most important to date. Two outcrops of the São Fidelis zone, Saudade and Sao Benedito, are now being exploited. Mining has been carried out, so far, on a small scale. The mineralization consists of a vein of nearly massive graphite in a very compact, partially granitized gneiss, with vertical dip. The vein is most irregular, with nests and lenses irregularly distributed along the fracture plane. The ore consists chiefly (particularly in Sao Benedito) of massive graphite appearing and disappearing along the main direction of the vein, forming a succession of irregular lenses in a sort of rosary design. The average width is more or less one foot, with variations from zero to 6 feet. The lode minerals are quartz, a little feldspar and garnet (the latter having its origin perhaps in the embedding gneisses), with rare ferruginous clays on the outcrops. In Saudade a little pyrite, pyrrhotite and chalcopyrite is found occasionally, as well as a carbonaceous mineral, very light, black and shiny, with conchoidal fracture, looking like a sort of jet. The composition of this mineral has not yet been established, but it seems to be very complex. The mineralization seems to be hydro-thermal at a rather low temperature, perhaps mesothermal. The paragenesis, according to samples from Saudade, seems to be as follows, the earlier minerals being listed first: pyrrhotite, pyrite, chalcopyrite, with quartz concomitant to the three and graphite formed later. Age of the mineralization is still uncertain, owing to the very in-

Jan 1, 1948

By J. J. McKetta, O. L. Culberson

INTRODUCTION The equilibrium constants for methane and for water, and for ethane and water have been calculated from experimental data for the two binary systems.2,3,11,12 These constants are for the two-phase systems for the temperature range of 100" to 340°F and the pressure range of 200 to 10,000 psia. The constants for ethane are greater than those for methane. Equilibrium constants for a natural gas from a natural gas-water system6 are about the same as those for methane. Equilibrium constants for water in all three systems are very nearly the same over the greater part of the range of temperatures and pressures studied. In ideal solutions, the vapor-liquid equilibrium constants for a component should be the same in all such solutions at a given temperature and pressure.5 Few solutions are ideal, and consequently the use of generalized constants often leads to error in situations where they do not actually apply. In decidedly non-ideal systems, it is desirable to have equilibrium constant data for the specific system. Vaporization equilibrium constants may be calculated either by thermodynamic methods or from experimental vapor-liquid equilibrium data. In the thermodynamic calculation. the equilibrium constant K may be shown to be" K = fP,/fl&apos;..........(1) where fPi is the fugacity of the pure component i at the given temperature and at the vapor pressure of the pure component, and fp is the fugacity of pure i in the gaseous state at the temperature and total pressure of the system. At best the thermodynamic calculation results in a K applicable to ideal solutions. Equilibrium constants calculated from Equation (1) have been found to have only limited agreement with experimental values, even though the solutions for which they were calculated approached ideality."&apos;" If the more volatile component is at a temperature above its critical, the accuracy of the equilibrium constant for the component is made more doubtful by the use of the vapor pressure term under such conditions. In order to make the thermodynamic calculation on a component above its critical temperature, one must resort to one of the empirical methods of extrapolation of the vapor pressure to conditions above its critical point.13 Application of the thermodynamic equilibrium constant then to the methane-water or the ethane-water systems would be risky since the systems are non-ideal, and the methane and ethane are above their critical in most petroleum operations. ETHANE-WATER SYSTEM In the temperature range of from 100° to 340°F and for pressures up to 10,000 psia, the vapor-liquid equilibrium data &apos;References given at end of paper. Manuscript received in the office of the Petroleum Branch March 16, 1951.

Jan 1, 1951