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By D. R. Conlon, E. R. Brownscombe

Errors in measurement of reservoir pressure include: (I) gauge errors, and (a) interpretation errors. Gauge errors may be reduced by: (a) reading charts with a comparator microscope, (b) use of hard metal stylus points to give clear lines, (6) insuring temperature equilibrium, and (d) making multiple tests. Check calibrations are equally as important as check well tests. A statistical analysis of such check runs shows quantitatively the advantages of a standardized "zero pressure line'' and of the use of field calibrations made at bottom-hole temperature on the same chart run in the well. Interpretation errors include: (a) fluid heads, which should be corrected to the top of the perforations on the basis of well-fluid gravity and from the top of the perforation to the datum level on the basis of reservoir-fluid gravity, and (b) gradients around the well. Conventional shut-in periods may be entirely inadequate to overcome these, necessitating a reservoir study based on build-up curve data, reservoir history, core analysis and any other available information. Interest in engineering studies of the pressure changes in oil and gas reservoirs is increasing. This type of work yields important information and doubtless will be greatly extended in the future. One of the factors limiting the work, however, is the accuracy with which the reservoir pressures can be determined; for if the inaccuracies in the pressure measurements represent an appreciable percentage of the pressure changes, the engineer is seriously handicapped, and must delay the reservoir study until the reservoir pressure has changed a relatively large amount. This delay can greatly reduce the benefits of such studies, at least from the point of view of aiding the leasing, drilling and completion programs. Thus, from the point of view of reservoir analysis, it is frequently important to obtain the most accurate pressure data possible early in the history of any field—particularly for the gas-distillate fields now being developed. This paper deals with improvements in the measurement of bottom-hole pressures that permit pressures even as high as 4000 or 5000 lb. per sq. in. to be determined with gauge errors of only+ ±I to 2 lb. per sq. in. Since pressures out in the reservoir rather than in the bottom of the hole concern the reservoir analyst, differences between bottom-hole pressures and actual reservoir pressures are also considered. In general, errors in pressure measurement may be divided into: 1. Gauge errors: a. Errors due to inaccuracies in reading the pressure charts. b. Errors due to irreproducibility of the pressure element and recording system. c. Errors caused by failure to allow the gauge to reach temperature equilibrium. 2. Internretation errors: a. Errors caused by a fluid head between the gauge position in tubing and the pressure at the datum level in the reservoir. b. Pressure gradients in the reservoir.

Jan 1, 1946

By D. R. Conlon, E. R. Brownscombe

Errors in measurement of reservoir pressure include: (I) gauge errors, and (a) interpretation errors. Gauge errors may be reduced by: (a) reading charts with a comparator microscope, (b) use of hard metal stylus points to give clear lines, (6) insuring temperature equilibrium, and (d) making multiple tests. Check calibrations are equally as important as check well tests. A statistical analysis of such check runs shows quantitatively the advantages of a standardized "zero pressure line'' and of the use of field calibrations made at bottom-hole temperature on the same chart run in the well. Interpretation errors include: (a) fluid heads, which should be corrected to the top of the perforations on the basis of well-fluid gravity and from the top of the perforation to the datum level on the basis of reservoir-fluid gravity, and (b) gradients around the well. Conventional shut-in periods may be entirely inadequate to overcome these, necessitating a reservoir study based on build-up curve data, reservoir history, core analysis and any other available information. Interest in engineering studies of the pressure changes in oil and gas reservoirs is increasing. This type of work yields important information and doubtless will be greatly extended in the future. One of the factors limiting the work, however, is the accuracy with which the reservoir pressures can be determined; for if the inaccuracies in the pressure measurements represent an appreciable percentage of the pressure changes, the engineer is seriously handicapped, and must delay the reservoir study until the reservoir pressure has changed a relatively large amount. This delay can greatly reduce the benefits of such studies, at least from the point of view of aiding the leasing, drilling and completion programs. Thus, from the point of view of reservoir analysis, it is frequently important to obtain the most accurate pressure data possible early in the history of any field—particularly for the gas-distillate fields now being developed. This paper deals with improvements in the measurement of bottom-hole pressures that permit pressures even as high as 4000 or 5000 lb. per sq. in. to be determined with gauge errors of only+ ±I to 2 lb. per sq. in. Since pressures out in the reservoir rather than in the bottom of the hole concern the reservoir analyst, differences between bottom-hole pressures and actual reservoir pressures are also considered. In general, errors in pressure measurement may be divided into: 1. Gauge errors: a. Errors due to inaccuracies in reading the pressure charts. b. Errors due to irreproducibility of the pressure element and recording system. c. Errors caused by failure to allow the gauge to reach temperature equilibrium. 2. Internretation errors: a. Errors caused by a fluid head between the gauge position in tubing and the pressure at the datum level in the reservoir. b. Pressure gradients in the reservoir.

Jan 1, 1946

By S. W. Ross

In 1936 a description of the plant (Fig 1) and process employed by the Electrolytic Zinc Co. of Australasia Ltd. for the recovery of zinc from zinc concentrate by the electrolytic process was prepared. † During the twelve years which have elapsed since the preparation of the earlier paper, several major changes in the metallurgy of the process have been introduced. It is the purpose of the present paper to give a general description of these changes and thus to bring up to date the description of the plant and process. Summary The major changes in Risdon practice since 1936 have been: 1. Replacement of two stage roasting by a preliminary roast followed by the flotation of all the leach residue and the roasting of the flotation concentrate. 2. Screening of all calcine fed to the pachucas. 3. Continuous leaching of calcine and improved classification of pachuca discharge. 4. Close control of hydrogen ion concentration during purification for iron removal. 5. Recovery of cobalt as a good grade oxide. 6. Production of part of the zinc output in the form of "four nines" metal (99.99 pct purity). 7. Closer spacing of electrodes thus increasing the potential output of cathode zinc per cell by 50 pct. Changes which are in prospect and for which construction work is proceeding at the present time involve the starting up of: 1. Two suspension roasters. 2. A contact acid plant to produce 150 tons‡ of acid per day and replacing the existing Mills Packard chamber plant. 3. Extra power station capacity to permit greater current flow to existing cell room units. This will increase the output of cathode zinc from about 245 to 290 tons per day. Plans for the future envisage the building of an ammonium sulphate plant, the first unit of which will produce about 50,000 tons per year, and improved treatment of zinc plant residue for the recovery of zinc, lead and other metals. At the end of this paper tables of metallurgical data are presented relating to the year ending June 30, 1948. Details of Changed Practice Output In 1936 the production of cathode zinc amounted to about 200 tons per day. This has since been increased to about 245 tons per day while plant extensions are practically complete which will permit of an output of about 290 tons per day in the near future. Roasting Division ROASTING POLICY A major change has occurred in the roasting policy. Twelve years ago the method in use was to carry out a two stage roast in the first stage of which sulphide sulphur was reduced to about 6 pct. The pre-roast calcine was re-roasted in modified Leggo furnaces using coal as fuel, sulphide sulphur being reduced to about 0.8 pct. The whole procedure was described on pp. 482491 of the earlier paper. Although this roasting procedure had certain advantages it possessed some distinct disadvantages. For instance, it appeared uneconomical to heat up the entire input of pre-roast calcine to roasting temperature by the expenditure of fuel in order to oxidise a few percent of sulphide sulphur. It was argued that if the pre-roast calcine were leached and a process could be developed for the recovery of a zinc sulphide concentrate from the leach residue, this concentrate, small in weight compared with the pre-roost calcine, would probably roast autogenously, thus virtually eliminating the expenditure of fuel as well as greatly increasing the weight of sulphur oxidised per square foot of furnace hearth area. The obvious method of producing a suitable concentrate from leach residue was by flotation. It will be recalled (see p. 495 of the earlier paper) that when two-stage roasting was practised the leach residue was classified, the granular fraction was ground and floated while the slime fraction was thickened, filtered and dried ready for shipment to a lead smelter. This process worked quite successfully. However, when trials were made of leaching a calcine carrying several percent of sulphide sulphur the granular fraction still floated well, but the slime fraction carrying 8-10 pct sulphide sulphur yielded very poor results when subjected to flotation. This fact held up the application of "pre-roast" leaching for many years. However, successful flotation of the slime fraction of leach residue was finally achieved and in August 1940 the slime flotation plant began operation, while the leach-

Jan 1, 1950

By J. H. Westbrook

IT is always of interest to document the origin of real microstructures which closely approach various idealized morphologies. Such instances constitute excellent bases for experimental studies of the effect of microstructure on properties. Typical examples include the eutectoid steels (lamellae) used by many investigators and the two phase Cu-Cr alloys (spheres) recently employed by Hibbard and Hart The present note describes a structure with cylindrical particles. In a study of certain alloys in the Cr-C system, an alloy was arc melted whose composition was very close to the aCr-Cr,C, eutectic (3.2 wt pct C, according to the latest phase diagram published by Bloom and Grant2). The alloy was not analyzed but previous experience with the melting technique employed indicates good correspondence between intended and analyzed compositions with negligible tungsten contamination from the electrode. The phases present in the alloy were identified by X-ray powder patterns as a chromium and Cr,C,. Typical micrographs, presented in Fig. 1, show unmistakable evidence of the cylindrical geometry of the carbide phase. The eutectic grains are randomly oriented and, consequently, only a few show the true fibrous nature of the carbide particles, most being cut so as to show slightly elliptical cross sections. Within any particular grain, remarkable uniformity is to be observed with respect to the size, shape, spacing, and orientation of the carbide particles. Length to diameter ratios have been estimated from the micrographs to be at least 250:1, with diameters of about lp. This alloy may be of further interest from different viewpoints. On the one hand, it might be used to check the applicability of the equations derived by Fullman3 for computing true particle dimensions of cylindrical particles from opaque random sections. On the other hand, it might be possible to extract the carbide particles from this alloy and repack them in a ductile matrix to form a synthetic structure of the type discussed by Metcalfe et al.4 With the large l:d ratio and the small diameter of the carbide fibers, it is likely that the interdefect distance in each fiber is sufficiently great to permit realization of a large fraction of the presumed high strength of the carbide. References 1W. R. Hibbard, Jr. and E. W. Hart: Dispersion Hardening of Copper-Chromium Alloys. AIME Trans., 1955, vol. 203, p. 200; Joun-nal of Metals, January 1955. 2D. S. Bloom and N. J. Grant: The System Chromium-Carbon. AIME Trans.. 1950. vol. 188. p. 41; JouRnal of Metals, January 1950. aR. L. Fullman: Measurement of Approximately Cylindrical Particles in Ooaaue Samples. AIME Trans.. 1953, vol. 197, p. 1267; Journal of Metals. ~ebtember 1953 4 A. G. Metcalfe, C. H. Sump, and W. C. Troy: Fiber Metallurgy. Metal Progress, 1955, vol. 67, p. 81.

Jan 1, 1958

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. W. III 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 Hubbert.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: (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, 1941

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 John R. Green

ROTARY kilns, varying in construction, are used in a variety of processes. Products differ and operating conditions vary according to economic requirements. All of these variables influence the degree of instrumentation and type of control necessary or profitable to install. Products fired in rotary kilns are the results of combined temperature and time treatment and the interrelation of these two factors should be kept constantly in mind in the perusal of this paper. MEASUREMENTS OF KILN TEMPERATURE The securing of proper temperatures within the rotary kiln, either of the charge or of the kiln lining in the high fire zone, are basic objectives of kiln operation; therefore the measurement of these high-temperature conditions merits first consideration.. For these measurements, radiation pyrometers, properly designed and protected, have proved adequate, particularly where records or automatic control are desired, while optical pyrometers may serve as portable indicators. Light-sensitive devices or photoelectric cells in this service are capable of producing records as well as certain forms of automatic control. High-temperature measuring devices for rotary-kiln work should have the measuring unit protected from dust, as a continued accumulation of deposit would lower the readings. Fig. I shows a typical application of radiation equipment to a rotary kiln, mounted far enough away from the sighting hole through the kiln hood to be protected from puffs or stingouts and with a sighting tube having a series of settling chambers, to prevent dust from reaching the lens. Although dust conditions inside the kiln itself usually fluctuate considerably and have their effect upon the readings at that moment, such effects are not cumulative and show as a continuous fluctuation in the temperature indicated, provided the radiation unit and recorder are responsive to such slight and rapid temperature fluctuations. The ambient temperature to which a measuring unit is subjected fluctuates from possibly o° to over 200°F., therefore these changes should have no effect upon the final reading or upon the life of the measuring unit. A very narrow sighting angle is desirable, so that measurement of temperature may be made as far as possible into the high fire zone without. effect of the flame temperature. There should be some sighting means on the measuring unit, so that the proper field may be viewed. Measurement of wall temperature taken at a point where the charge has just left the wall is less affected by dusting conditions than when taken elsewhere. Factors Influencing Temperature Table I gives an outline of the factors that produce kiln temperatures and com¬ments on the possibility of measurement and control. Of greater number and of varied influ¬ence are the factors that control when, where and how the heat is delivered to the kiln, the type of flame and the rate of

Jan 1, 1941

By I. W. Walling, F. B. Plummer

Bastin and others have demonstrated the presence of sulphate-reducing bacteria in oil wells producing salt water. Analyses show that at 125°F. bacteria alone reduce sulphates in East Texas salt water by only 3 per cent in 30 days, forming principally hydrogen sulphide and sodium or calcium hydroxide. The hydrogen sulphide reacts with soluble iron compounds to produce iron sulphide in the absence of oxygen, and ferric hydroxide where oxygen and air are present. Where iron or other metals and iron sulphide are already in the water containing bacteria, I00 per cent reduction of the sulphates is complete in 30 days or less; where magnesium is present, reduction is complete in 15 days. The effects of changes in temperature, of different bactericides, and of concentration of bactericides on sulphate reduction are described in detail and illustrated by tables and curves. The significance of sulphate reduction on corrosive properties of oil-field water is briefly discussed. Chemistry of East Texas Oil-field Water The chemical composition of the water in the East Texas field salt-water disposal pits is shown in Table I. This salt water corrodes pipes, pumps, and steel tanks excessively, principally because of its content of oxygen, hydrogen, sulphide, and carbonic acid. The carbonates in the water when exposed to air tend to give up carbon dioxide and to precipitate calcium carbonate in large quantities. This forms coatings in tanks, settling pits, and ditches, and soon, unless removed by acid treatment, will effectively clog an injection well and reduce the flow of water back into the ground. The iron compounds in the water, most of which are dissolved from tanks and pipes, are in the ferrous form, and when exposed to the air are oxidized from soluble ferrous carbonate to insoluble ferric hydroxide, resulting in the precipitation of a gelatinous precipitate that will clog oil-field sands effectively: 2Fe(HCO3)2 + O + H2O -> 2Fe(OH)3 + 4CO2 The sulphates in strong sulphate-bearing water may be attacked by swarms of sulphate-reducing bacteria and more hydrogen sulphide may be produced. The hydrogen sulphide reacts with soluble minerals in the water to form black sludges. A knowledge of the chemistry of each reaction in oil-field water is important to the oil-field engineer, since means must be devised to prevent the formation of harmful precipitates that clog injection wells and to prevent corrosion of tanks by harmful products of biochemical processes. Methods of Treating Water before Injection into Ground* The open system of handling the waste salt water in the East Texas field consists of aerating and chemically treating the water to precipitate the unstable iron,

Jan 1, 1946

By F. B. Plummer, I. W. Walling

Bastin and others have demonstrated the presence of sulphate-reducing bacteria in oil wells producing salt water. Analyses show that at 125°F. bacteria alone reduce sulphates in East Texas salt water by only 3 per cent in 30 days, forming principally hydrogen sulphide and sodium or calcium hydroxide. The hydrogen sulphide reacts with soluble iron compounds to produce iron sulphide in the absence of oxygen, and ferric hydroxide where oxygen and air are present. Where iron or other metals and iron sulphide are already in the water containing bacteria, I00 per cent reduction of the sulphates is complete in 30 days or less; where magnesium is present, reduction is complete in 15 days. The effects of changes in temperature, of different bactericides, and of concentration of bactericides on sulphate reduction are described in detail and illustrated by tables and curves. The significance of sulphate reduction on corrosive properties of oil-field water is briefly discussed. Chemistry of East Texas Oil-field Water The chemical composition of the water in the East Texas field salt-water disposal pits is shown in Table I. This salt water corrodes pipes, pumps, and steel tanks excessively, principally because of its content of oxygen, hydrogen, sulphide, and carbonic acid. The carbonates in the water when exposed to air tend to give up carbon dioxide and to precipitate calcium carbonate in large quantities. This forms coatings in tanks, settling pits, and ditches, and soon, unless removed by acid treatment, will effectively clog an injection well and reduce the flow of water back into the ground. The iron compounds in the water, most of which are dissolved from tanks and pipes, are in the ferrous form, and when exposed to the air are oxidized from soluble ferrous carbonate to insoluble ferric hydroxide, resulting in the precipitation of a gelatinous precipitate that will clog oil-field sands effectively: 2Fe(HCO3)2 + O + H2O -> 2Fe(OH)3 + 4CO2 The sulphates in strong sulphate-bearing water may be attacked by swarms of sulphate-reducing bacteria and more hydrogen sulphide may be produced. The hydrogen sulphide reacts with soluble minerals in the water to form black sludges. A knowledge of the chemistry of each reaction in oil-field water is important to the oil-field engineer, since means must be devised to prevent the formation of harmful precipitates that clog injection wells and to prevent corrosion of tanks by harmful products of biochemical processes. Methods of Treating Water before Injection into Ground* The open system of handling the waste salt water in the East Texas field consists of aerating and chemically treating the water to precipitate the unstable iron,

Jan 1, 1946

Lead-zinc production in Canada accounted for 19% of the total value of metals and minerals produced in 1976, says Keith C. Hendrick, president of Noranda Sides Corp. Mine production of recoverable zinc w as 1. 158 million toils of 27.4% of free world production in 1976. Slab zinc pro¬duction was placed at 472,000 tons that year. Canadian mine production of recoverable lead was 247,000 tons or 10.6% of the free world production. Refined lead that year was at 177,000 tons.

Jan 11, 1977

By H. R. Stahl

The disseminated lead district of Southeast Missouri lies 70 miles south of St. Louis. The only metal of economic importance in the ore is lead, but minor amounts occur of iron, zinc, copper, cobalt, nickel, silver and cadmium. These minor metals, as might be expected, tend to concentrate in both the flotation and table middlings. In the past, some zinc has been recovered by differential flotation. The silver and cadmium tend to associate with the zinc, and some recovery is made of these two metals at the smelter. The lead occurs as galena, or lead sulphide. The minor metals are also present as sulphides. The presence of oxidized lead is suspected, but has not been definitely proved. The gangue is chiefly dolomitic limestone, with small amounts of sandstone and shaly glauconitic and chloritic minerals. The ore contains about 3.25 per cent lead, occurring as galena disseminated through the gangue. Some of the galena is very finely disseminated, and the mineral in the gravity tailings, or chat; is seldom or never in the free state. The method of treatment of the ore has been evolved over a period of 75 years; hence probably is not far from correct. The milling processes have tended toward simplification during the past 15 years, but recoveries have never been sacrificed for the sake of simplicity. In many cases, simplification of flowsheets has led to better recoveries and operating economies. A simplified flowsheet is shown in Fig. I. The St. Joseph Lead Co. operates four mills in the Flat River district, with a normal total capacity of about 21,000 tons per day, divided as follows: Federal mill, 11,000 tons; Leadwood mill, 4500 tons; Desloge mill, 3400 tons and Bonne Terre mill 2400 tons. The methods of concentration used at these plants are similar in principle; they differ mainly in details and in the type and size of the equipment used. The process of concentration may be divided into five main steps; i.e., dry crushing, wet grinding, classification with its attendant desliming, tabling and flotation. A brief description of these processes will be given, noting important differences in the various plants, and also noting the more important changes or developments of the past few years. Dry Crushing The run-of-mine ore is hoisted in skips, the capacity of these skips ranging from 23/4 tons to 81/4 tons. At the Desloge plant, the ore is hoisted in 2-ton mine cars on self-dumping cages. All ore is hoisted at shafts adjacent to the various crushing plants, no ore being transported on the surface. Primary breaking at all the mills is accomplished with 9E Telsmith breakers. Preliminary screening is provided ahead of most of these crushers, either by station-

Jan 1, 1943

By H. R. Stahl

The disseminated lead district of Southeast Missouri lies 70 miles south of St. Louis. The only metal of economic importance in the ore is lead, but minor amounts occur of iron, zinc, copper, cobalt, nickel, silver and cadmium. These minor metals, as might be expected, tend to concentrate in both the flotation and table middlings. In the past, some zinc has been recovered by differential flotation. The silver and cadmium tend to associate with the zinc, and some recovery is made of these two metals at the smelter. The lead occurs as galena, or lead sulphide. The minor metals are also present as sulphides. The presence of oxidized lead is suspected, but has not been definitely proved. The gangue is chiefly dolomitic limestone, with small amounts of sandstone and shaly glauconitic and chloritic minerals. The ore contains about 3.25 per cent lead, occurring as galena disseminated through the gangue. Some of the galena is very finely disseminated, and the mineral in the gravity tailings, or chat; is seldom or never in the free state. The method of treatment of the ore has been evolved over a period of 75 years; hence probably is not far from correct. The milling processes have tended toward simplification during the past 15 years, but recoveries have never been sacrificed for the sake of simplicity. In many cases, simplification of flowsheets has led to better recoveries and operating economies. A simplified flowsheet is shown in Fig. I. The St. Joseph Lead Co. operates four mills in the Flat River district, with a normal total capacity of about 21,000 tons per day, divided as follows: Federal mill, 11,000 tons; Leadwood mill, 4500 tons; Desloge mill, 3400 tons and Bonne Terre mill 2400 tons. The methods of concentration used at these plants are similar in principle; they differ mainly in details and in the type and size of the equipment used. The process of concentration may be divided into five main steps; i.e., dry crushing, wet grinding, classification with its attendant desliming, tabling and flotation. A brief description of these processes will be given, noting important differences in the various plants, and also noting the more important changes or developments of the past few years. Dry Crushing The run-of-mine ore is hoisted in skips, the capacity of these skips ranging from 23/4 tons to 81/4 tons. At the Desloge plant, the ore is hoisted in 2-ton mine cars on self-dumping cages. All ore is hoisted at shafts adjacent to the various crushing plants, no ore being transported on the surface. Primary breaking at all the mills is accomplished with 9E Telsmith breakers. Preliminary screening is provided ahead of most of these crushers, either by station-

Jan 1, 1943

By H. R. Stahl

THE disseminated lead district of Southeast Missouri lies 70 miles south of St. Louis. The only metal of economic importance in the ore is lead, but minor amounts occur of iron, zinc, copper, cobalt, nickel, silver and cadmium. These minor metals, as might be expected, tend to concentrate in both the flotation and table middlings. In the past, some zinc has been recovered by differential flotation. The silver and cadmium tend to associate with the zinc, and some recovery is made of these two metals at the smelter. The lead occurs as galena, or lead sulphide. The minor metals are also present as sulphides. The presence of oxidized lead is suspected, but has not been definitely proved. The gangue is chiefly dolomitic limestone, with small amounts of sandstone and shaly glauconitic and chloritic minerals. The ore contains about 3.25 per cent lead, occurring as galena disseminated through the gangue. Some of the galena is very finely disseminated, and the mineral in the gravity tailings, or chat, is seldom or never in the free state. The method of treatment of the ore has been evolved over a period of 75 years; hence probably is not far from correct. The milling processes have tended toward simplification during the past 15 years, but recoveries have never been sacrificed for the sake of simplicity. In many cases, simplification of flowsheets has led to better recoveries and operating economies. A simplified flowsheet is shown in Fig. I. The St. Joseph Lead Co. operates four mills in the Flat River district, with a normal total capacity of about 21,000 tons per day, divided as follows: Federal mill, 11,000 tons; Leadwood mill, 4500 tons; Desloge mill, 3400 tons and Bonne Terre mill 2400 tons. The methods of concentration used at these plants are similar in principle; they differ mainly in details and in the type and size of the equipment used. The process of concentration may be divided into five main steps; i.e., dry crushing, wet grinding, classification with its attendant desliming, tabling and flotation. A brief description of these processes will be given, noting important differences in the various plants, and also noting the more important changes or developments of the past few years. DRY CRUSHING The run-of-mine ore is hoisted in skips, the capacity of these skips ranging from 2 ¾ tons to 8 ¼ tons. At the Desloge plant, the ore is hoisted in 2-ton mine cars on self-dumping cages. All ore is hoisted at shafts adjacent to the various crushing plants, no ore being transported on the surface. Primary breaking at all the mills is accomplished with 9E Telsmith breakers. Preliminary screening is provided ahead of most of these crushers, either by station-

Jan 1, 1943

By J. K. Stanley

IT has been known for many years that the magnetic properties of a ferromagnetic material are very sensitive to internal strain. Any structure-sensitive property such as ferromagnetism, which is a-function of the regularity of the atomic arrangement, and discontinuities such as grain boundaries must show deviations from an equilibrium state when the regularity of the lattice is disturbed. Experience has shown that such distortions have a marked effect on the magnetic properties of a ferromagnetic material. The purpose of this paper is to show the feasibility of using magnetic methods in detecting internal strains. More specifically, it will be shown how such properties as permeability, remanence, and coercive force change on cold- working of aluminum iron and how these magnetic properties change during the annealing below the recrystallization temperatures. This work was conducted for two reasons. One was the desire to see in what manner strains are relieved in aluminum iron at low temperatures-a subject of practical importance in soft magnetic materials-and the other was the feeling that a study of cold-working and strain relief might shed some light honour understanding of what takes place in the deformation of metals DEFINITION IN TERMS It is necessary to define the terminology used in this paper because magnetic nomenclature is usually unfamiliar to the metallurgist, and certain other terms, such as recovery, have been used ambiguously by metallurgists and others. Magnetic terms are best explained by reference to the common hysteresis loop of ferromagnetic materials (Fig. I). In Fig. I the coordinates are magnetic induction, B, (ordinate) and magnetizing force, H, (abscissa). The c.g.s. unit for the magnetic induction is the gauss, the number of flux lines per square centimeter, and the unit for the magnetizing force is the oersted, the number of lines of force per centimeter in air or vacuum. If one starts with a completely de- magnetized specimen and magnetizes it, a typical virgin magnetization curve beginning at the origin is obtained. This curve is also the locus of tips of a family of smaller hysteresis loops. From this curve it is to be noted that the induction is not proportional to the magnetizing force, and that it appears to approach a limiting induction called the saturation value, and does so for values of H of 1000 oersteds. The ratio of the magnetic induction (gauss) to the magnetizing force (oersteds) is called permeability, p. Thus µ = B/H and if the magnetizing force H (oersteds) is specified as 10 or IOO oersteds, the permeability is written as µ10 or µ100. This is the accepted notation, and will be used in this paper. The values of permeability for H = 10 or 100 are chosen as indices for

Jan 1, 1945

By R. T. Webber, R. B. Owen

The vapor phase deposition of molybdenum and tungsten from the hexacar-bonyls has been thoroughly investigated by Lander and Germerl, and shown to yield well-bonded coatings on a variety of hot metallic surfaces. They also demonstrated that chromium coatings2'3 could be similarly deposited, but did not investigate this metal extensively. Recently, conditions have been described by which this compound can be prepared in quantities sufficient for laboratory scale plating experiments, and it is the purpose of the present communication to outline the results of preliminary experiments upon the plating of chromium from its carbonyl. Apparatus and Technique The apparatus used in these experiments is shown in Fig I. At one end, connected by a short length of rubber tubing, is a Cenco Hypervac 20 pumping continuously. At the opposite end is a needle valve which allows a small stream of purified hydrogen to flow through the apparatus. Connected in series are a trap T1, maintained at the temperature of dry ice and acetone, the plating chamber, and two traps T² and T3 containing crystals of the carbonyl. T² is kept in a constant temperature bath (not shown), while T³ is at room temperature. Stopcock S6 allows the hydrogen flow to by-pass the carbonyl traps. A McLeod gauge is connected so that it can measure the pressure at the pump through Se, or the pressure at the plating chamber through S9 and S6. The standard (12 liter) volume connected by Sb is used to determine the volume included by S4, S6, SB, Slot the needle valve and the McLeod gauge. Knowledge of this volume allows calibration of the flow of hydrogen. The apparatus* is built entirely of pyrex, except for the needle valve, the rubber connection to the Hypervac, and the cover of the plating chamber. The plating chamber is water-jacketed, with a large brass top allowing insertion of the specimen to be plated. Set into the brass top is a ground glass stopper through which thermocouple leads are sealed. The thermocouple wires (pure Pt, and 87 pct Pt-13 pct Rh) serve the double purpose of measuring the specimen temperature and suspending it in the chamber. Surrounding the chamber is the coil of an induction heater, H. The apparatus is designed to allow variation of the following conditions: (I) the specimen temperature, from room temperature to about 1000°C, (2) the carbonyl vapor pressure, from 40 to 220 microns, and (3) the hydrogen flow. In several runs, 0.8 to 1.8 PCt of hydrogen sulfide was introuced through a needle valve that replaced S4. The vapor pressure of the carbonyl is

Jan 1, 1949

By L. G. Graper, E. L. Ramsey

THE present program of increased production of armament and lend-lease material for mechanized war has created a problem for the shops that must do the machining. They have naturally turned to the steel producers to provide them with a steel that can be accurately machined with the greatest possible speed. The adding of sulphur to steel as an aid to machinability has long been practiced; however, it may cause red shortness in rolling and may result in billets with seams and cracks, which must be removed before rolling can be continued. This conditioning of billets slows down production and adds to the cost of making steel. Experience with various grades of sulphur steel made to different specifications has shown that Bessemer sulphur steels roll better than similar grades made in the open hearth, and also that low-carbon steels that are open or semi-killed as tapped roll better than completely deoxidized or killed steels. This has led to the belief that oxygen plays an important part in the rolling qualities of sulphur-bearing steels. It is generally held that the red shortness of sulphur steels is due to the low melting point of the sulphide compounds that separate out at the grain boundaries when the steel solidifies. At the rolling temperature of steel these sulphide compounds melt and cause cracking. ADDING SULPHUR AS A SULPHITE These difficulties can be avoided if sulphur is introduced into the steel by means of an anhydrous sulphite, such as sodium sulphite or sodium bisulphite, which-upon decomposition-forms sulphur dioxide. It is assumed that, in this case, the oxygen of the sulphur dioxide forms part of the sulphide inclusions in the steel, and raises their solidification temperature enough so that the formation in the steel of a low-melting eutectic is prevented. The result is the prevention of hot shortness and a more uniform distribution of inclusions. Fig, I shows photomicrographs of steels made with sulphur, while Fig, 2 illustrates the structure of steels made with sodium bisulphite. These samples were cut from the steels as cast-that is, before rolling and were polished and lightly etched to bring out the grain boundaries. Fig. I shows that sulphide inclusions segregate more or less at the grain boundaries, and it can be seen how the continuity of these low-melting inclusions would cause red shortness. Fig. 2 shows that the sulphide inclusions are more or less uniformly distributed throughout the matrix. This arrangement of inclusions has no apparent effect on the rolling qualities of a steel. This statement has been substantiated by the observation of the rolling of over 6000 tons of steel made with a sulphite addition. RESULTS As stated before, the sulphur dioxide is added to steel by means of the decompcsi-

Jan 1, 1942

By S. F. Urban, G. Derge

At the time this investigation was initiated variations were observed in the quality of different heats of basic electric-furnace steels, although they had been made under purportedly similar conditions. This indicated that these variations were related to normally unobserved variations in the character of the bath, especially during the "refining" period. Although the slags used for this operation are commonly referred to as "reducing" slags, and are regarded as having a "deoxidizing" action on the bath, no reliable measurements of this action were available. It was decided, therefore, to study the behavior of oxygen in the metal for ' large enough number of heats to learn its normal variations and to relate these variations to regularly observed conditions of operation. The most important general conclusion reached was that the time a heat is held under a reducing 'lag should be reduced to a minimum, which was about an hour and a half under the conditions studied. Beyond this time no observable beneficial effects are obtained from the reducing slags, insofar as oxygen and sulphur content of the bath are concerned. It is likely that added time may increase hydrogen content as well as heat costs. The oxygen content of the bath is never reduced below the amount in equilibrium with the carbon. Concurrently with this work, J. S. Marsh1 presented data that supplement and support many of the results of this paper. Procedure The work was done in two different electric-furnace shops of the same plant. The selection of heats was made primarily with regard to the convenience of heat schedules. companion slag and metal samples were taken just prior to slag-ofi and throughout the refining period for an initial set of 17 heats, and before slag-off and near tap for an additional I3 heats. slag Sampling presented no particular problem; a spoonful of slag was poured onto a metal plate and when sufficiently cool was sealed in an airtight container. Sampling the metal for oxygen presented a more complex problem. Because of the lack of previous experience in sampling electric-furnace baths, a number of methods were compared. Chilled Wedge Samples2 An unkilled metal sample is removed from the furnace in a well-slagged car-bometer spoon and poured immediately into a heavy copper mold designed to provide rapid freezing. Three different types of molds were used to determine the optimum for electric-furnace hats. The Big wedge Mold (a) was a 4 by 4 by 4-in. split copper mold, yielding a wedge-shaped casting I in. wide, 1 1/2 in. long, and 34 in. maximum thickness, and was finally adopted as the most satisfactory design.

Jan 1, 1948

By S. F. Urban, G. Derge

At the time this investigation was initiated variations were observed in the quality of different heats of basic electric-furnace steels, although they had been made under purportedly similar conditions. This indicated that these variations were related to normally unobserved variations in the character of the bath, especially during the "refining" period. Although the slags used for this operation are commonly referred to as "reducing" slags, and are regarded as having a "deoxidizing" action on the bath, no reliable measurements of this action were available. It was decided, therefore, to study the behavior of oxygen in the metal for ' large enough number of heats to learn its normal variations and to relate these variations to regularly observed conditions of operation. The most important general conclusion reached was that the time a heat is held under a reducing 'lag should be reduced to a minimum, which was about an hour and a half under the conditions studied. Beyond this time no observable beneficial effects are obtained from the reducing slags, insofar as oxygen and sulphur content of the bath are concerned. It is likely that added time may increase hydrogen content as well as heat costs. The oxygen content of the bath is never reduced below the amount in equilibrium with the carbon. Concurrently with this work, J. S. Marsh1 presented data that supplement and support many of the results of this paper. Procedure The work was done in two different electric-furnace shops of the same plant. The selection of heats was made primarily with regard to the convenience of heat schedules. companion slag and metal samples were taken just prior to slag-ofi and throughout the refining period for an initial set of 17 heats, and before slag-off and near tap for an additional I3 heats. slag Sampling presented no particular problem; a spoonful of slag was poured onto a metal plate and when sufficiently cool was sealed in an airtight container. Sampling the metal for oxygen presented a more complex problem. Because of the lack of previous experience in sampling electric-furnace baths, a number of methods were compared. Chilled Wedge Samples2 An unkilled metal sample is removed from the furnace in a well-slagged car-bometer spoon and poured immediately into a heavy copper mold designed to provide rapid freezing. Three different types of molds were used to determine the optimum for electric-furnace hats. The Big wedge Mold (a) was a 4 by 4 by 4-in. split copper mold, yielding a wedge-shaped casting I in. wide, 1 1/2 in. long, and 34 in. maximum thickness, and was finally adopted as the most satisfactory design.

Jan 1, 1948

By L. A. Roe

The flotation of fine particles less than five microns in diameter and the flotation of liquids from liquids generally requires bubble generation methods different than methods used in dispersed air type machines used in mineral flotation. Conditions in the separatory zone in mineral cells are generally more dynamic as compared to the quiescent conditions often absolutely necessary for the flotation of fine particles and liquids. An example of quiescent flotation with very fine bubbles is the lithium flotation plant which operated at Trona, California for 34 years, until it ceased operations in 1978. A review of the process development and the commercial flotation operations at Trona is given in this paper. The clarification of oil field brines by flotation of the oil from the brine has been used for over twenty-five years. Recent installations related to the Prudhoe Bay oil field and the Alaska pipeline are described.

Jan 1, 1980