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By A. Simkovich, R. W. Lindsay

The possibility of using sodium sulfide slags to remove copper from ferrous alloys has been investigated by Jordan1 and by Langenberg.2, 3 In these studies, such slags were determined to be capable of removing copper and sulfur from the melt. The present work represents additional effort to clarify the effects of temperature on copper removal. The experiments were performed in a 17-lb induction furnace. Graphite crucibles contained the melts and kept the baths saturated with carbon. Temperatures were measured with a calibrated optical pyrometer and were controlled by manipulation of power input to the furnace. Estimated accuracy of temperatures in this investigation is ± 10°C (18°F) for measurements prior to slag additions, and + 20°C (36°F) after slag formation. The procedure consisted of melting 800 g of electrolytic iron. During this step, powdered graphite covered the exposed iron surface. After a predetermined temperature was reached, copper shot was added. A sample of the molten alloy for chemical analysis was then aspirated into a silica sheath. Next, a slag-forming mixture of sodium sulfite and graphite was added instantaneously to the melt. The sodium sulfite amounted to one-tenth the charge weight of iron; sufficient graphite was added to combine with oxygen in the sodium sulfite, assuming formation of carbon monoxide and reduction of the sulfite to sulfide. Subsequent to the slag addition, the molten alloy was sampled periodically, with the exception of heat A in which no intervening samples were taken between the slag addition and the end of the run. The iron was poured into a graphite mold, and the ingots sectioned and drilled for samples. Results of selected heats are presented in Table I. Analyses of samples drawn from the iron prior to slag addition are listed under zero time. Two samples from heat D were reported with copper contents greater than the initial concentration in the bath. Owing to the gradual but complete disappearance of slag during this heat, it is believed copper momentarily became more concentrated in the upper portion of the bath while reverting from the slag. This is the region from which samples were drawn. It should be noted that analysis of the ingot was equal to the copper content at the time of slag addition. The terminal temperatures of heats D and E, and the initial sulfur content of heat A are also to be noted. Because of the large temperature drop which occurred when slag was formed in heat D, power input to the furnace was increased in heat E after the slag addition, causing a higher terminal temperature. In heat A, the initial sulfur concentration was relatively high as compared to heats B through E owing to contamination by some slag remaining in the crucible from a previous heat. It is evident from Table I that copper was removed at the onset of slag formation. Roughly 30 pct of the copper was taken into the slag, with the exception of heat D, which had approximately 50 pct removed. For a comparatively short time of slag-metal contact, it appears that no gain is to be made in copper removal through use of high or low temperatures. If the slag initially formed remains in contact with the iron for an extended period, temperature has a marked effect upon copper removal, as can be seen by studying results for the two extremes in temperature. At about 1425°C, the copper level remained relatively constant after the initial removal by the slag. However, in the region of 1670°C, a definite reversion of copper occurred. Reversion was incomplete in heat D, and complete in heat E. The final temperatures of heats D and E differed by about 75°C. This temperature difference is thought to be the reason for only partial copper reversion in heat D. It is believed the effects of temperature noted above are related to the evolution of a white fume, which appeared in every run except heat A. (In the case of heat A, the fume was practically indiscernible.) After each slag addition, a yellow flame formed for about 5 sec. When the flame subsided, a white fume appeared. Upon contact with surrounding cooler surfaces, this fume deposited as a white solid. In the experiments made at 1425°C, evolution of fume continued unchanged to the end of the runs. However, heats D and E exhibited a different behavior. A very noticeable decrease in fume evolution from heat D was observed. Furthermore, this heat had much less slag remaining than did runs A through C when the experiments were terminated. No slag remained at the end of heat E; evolution of fume from this heat ceased prior to pouring. Spec-trographic analysis of the white deposit indicated sodium to be the major metallic element, with the maximum concentration of iron and copper as 0.1 and 0.01 pct, respectively. It is supposed the white fume observed in these experiments is principally sodium oxide (Na2O), formed by oxidation of sodium in the slag and subsequent sublimation. (Sodium oxide is a white to gray substance in the solid state; at 1275oC, it sublimes.4) According to this mechanism, elevated temperatures would accelerate removal of sodium from the slag, sulfur pickup by the

Jan 1, 1961

By R. H. Gautschi, F. C. Langenberg

Rare-earth additions were made to laboratory heats of Type 310 stainless to observe their effect on as-cast ingot structure, nitrogen and sulfur contents, and nonmetallic inclusions. Lanthanum had a grain-refining effect in 30-lb ingots, but results with 200-lb ingots were inconsistent. Cerium, lanthanum, and misch metal lowered the sulfur content when the sulfur exceeded 0.015 pct and the rare-earth addition was greater than 0.1 pct. The rare-eardh content in the metal dropped very rapidly within the first few minutes after the addition. The size, shape, and distribution of nonmetallic inclusions was not changed in 30-lb ingots, but changes were noticed in larger ingots. RARE-earth* additions have been made to austenitic Cr-Ni and Cr-Mn steels to improve their hot workability. The high alloy content of these steels often results in a considerable resistance to deformation and inherent hot shortness at rolling temperatures, particularly in larger ingots. Rare earths in the metallic, oxide, or halide form are usually added to the steel in the ladle after deoxidation although they can be added in the furnace prior to tap or in the molds during teeming. The literature- indicates that the effects of rare-earth treatments on these stainless steels are not consistent, and sometimes even contradictory. Since no mechanism has been presented which satisfactorily accounts for the claimed improvements, the effects of rare earths are a qualitative matter. The work described in this paper was initiated to expand the knowledge of the effects of rare-earth additions on melting variables such as ingot structure, chemical analysis, and nonmetallic inclusions. REVIEW OF LITERATURE Ingot Structure—Rare-earth additions to stainless steels have been reported to cause a change in primary ingot structure in that there are fewer coarse columnar grains. However, the results are inconsistent. While one investigation1 has shown a large reduction in coarse columnar crystals, another2 has been unable to observe this effect, particularly when small ingots were poured. Post and coworkers3 observed ingot structures for a number of years and found that the columnar type of structure is not definitely a cause of any particular trouble in rolling or hammering, provided the alloy is ductile. Knapp and Bolkcom4 found rare-earth additions to be quite effective in preventing grain coarsening in Type 310 stainless steel. Chemical Analysis—Many effects of rare-earth treatment on chemical analysis have been claimed in the literature. Russell5 observed that some sulfur is removed by rare-earth metals, and that a high initial sulfur content improved the efficiency of sulfur removal. Lillieqvist and Mickelson6 report that rare-earth treatment causes sulfur removal in basic open-hearth furnaces, but not in basic lined induction furnaces. Knapp and Bolkcom found no sulfur removal in acid open-hearth and acid electric furnaces, probably because the acid slag can not retain sul-fides. snellmann7 showed that sulfur could be lowered apprecfably with rare-earth additions; however, a sulfur reversion occurred with time. Langenberg and chipman8 studied the reaction CeS(s) = Ce(in Fe) + S(in Fe), and found the solubilit product [%Ce] [%S] equal to (1.5 + 0.5) X 10-3'at 1600°C. Results in 17 Cr-9 Ni stainless were about the same as those in iron. Beaver2 treated chromium-nickel steels with 0.3 pct misch metal and observed some reduction in the oxygen content. He also noted an inconsistent but beneficial effect of rare earths when tramp elements were present in amounts sufficient to cause difficulty in hot working. It is not known whether rare earths reduce the content of the tramp elements or change the form in which these elements appear in the final structure. No quantitative data are available concerning a possible effect of rare-earth treatment on hydrogen and nitrogen contents. However, Schwartzbart and sheehan9 stated that additions of rare earths had no effect on impact properties when the nitrogen content was low (0.006 pct), but served to counteract the adverse effects of high nitrogen content (0.030 pct) on these properties. Knapp and Bolkcom4 analyzed open-hearth heats in the treated and untreated conditions and found the nitrogen content (0.006 pct) to be unaffected. These two results lead to the speculation that rare-earth additions can reduce the nitrogen content to a certain level. Decker and coworkers10 have observed that small amounts of boron or zirconium, picked up from magnesia or zirconia crucibles, increased high-tem-

Jan 1, 1961

By C. G. Dunn, P. K. Koh

THE primary recrystallization texture in cold-rolled silicon iron, which is the matrix texture for developing the Goss texture or the cube-on-edge texture by secondary recrystallization at temperatures near 900°C,1-5,21 has not been adequately described. Thus from the present available information, it is not possible to explain satisfactorily both the grain-growth selectivity of the matrix and the often observed magnetic torque curve, which itself provides some information on the texture. Also lacking is information on the special annealing texture that develops in this material when the annealing temperature is 1200°C or higher and the rate of rise to temperature is extremely rapid. From the work of May and Turnbull5 and from unpublished work, it was known that isothermal annealing near 1200°C tends to reduce the extent of secondary recrystallization and that a much weaker cube-on-edge texture results if appreciable normal grain growth replaces secondary recrystallization. Koh and Dunn6,7 have obtained additional information on complex primary recrystallization textures from further studies made after normal grain growth. In these instances the initial textures were retained during normal grain growth. A similar result reasonably could be expected in the present study except for the presence of a grain growth ihibitor1-5,8,9.21 and its tendency to allow only a few grains to grow. However, any information on the orientations of grains in the special annealing texture, even if far from representative of the initial matrix texture, would provide useful information on the nature of the matrix texture. In the present texture study the method of Newkirk and Bruce,10 which is based on the methods of Geisler" and Schwartz, 12 is used to obtain a complete (110) pole figure of the primary recrystallization texture. The high-temperature annealing texture is determined simply from the orientations of a large number of selected grains. The kinetic nature of the process that produces the annealing texture is treated elsewhere13 and it is shown that a form of secondary recrystallization with a very high rate of nucleation occurs during rapid annealing at high temperatures. EXPERIMENTAL PROCEDURE Commercial 0.014-in. cold-rolled silicon iron strip (3.16 wt pct Si), prepared by two stages of cold rolling with an intermediate short anneal, was given a decarburizing 3-min anneal at 800°C. Me-tallographic studies indicated complete recrystallization. Short-time anneals at 900° C and at higher temperatures proved that secondary recrystallization had not begun at 800°C, in fact, the short additional anneals were still in the induction period of secondary recrystallization. A rapid rise to an annealing temperature of 1260°C (2300°F) was obtained in a BaC12and NaCl fused salt bath. The structures that resulted from anneals in the range 12- to 1000-sec duration were relatively finegrained, even though the growth was a form of exaggerated grain growth13 or secondary recrystallization with a high nucleation frequency.= Many of the grains were large enough for an X-ray study using a 5-mil X-ray beam. A transmission Laue Camera and an optical-mechanical stage for moving the grains into the X-ray beam were used. A total of 325 grains were X-rayed in this manner and the grain orientations determined. Complete (110) pole figures were obtained for the primary recrystallization texture using CoKa radiation at 30 kv in the back-reflection range, such as (220) for (110) poles, as described recently by Newkirk and Bruce.10 The low voltage serves to reduce the spurious white radiation to a minimum. A filter of 0.001-inch iron foil was located in front of the detector slit for transmission and in front of the beam slit for back reflection. A new and improved specimen holder extended the useful tilting angle range for transmission to 70 deg instead of 60 deg as previously reported.' A torque magnetometer was used to obtain the magnetic torque curves for a number of l-in. diam disk specimens. RESULTS The (110) pole figure of the material after recrystallization at 800°C is shown in Fig. 1. The positions of the pole concentrations are found to be

Jan 1, 1961

By C. S. Smith

QUANTITATIVE methods were used to good effect in the earliest days of metallography1-3 but they mysteriously passed into virtual disuse until the important paper4 by Howard and Cohen in 1947. Various devices for the measurements of integrated linear intercepts have been constructed, primarily by petrographers. The hydraulically operated stage designed by the writer5 was satisfactory, but prolonged use caused air bubbles to become entrapped, which introduced back-lash. A new stage has been constructed which avoids this defect and provides better integration when measuring directional structures. This utilizes a circular traverse, and integration is achieved by counting electrical pulses produced by a photocell and a light chopper.* The stage is shown diagramatically in Fig. 1 and in photograph, Fig. 2. It is of simple construction. Mounted on a plate with clamps to fit the stage of a standard metallurgical microscope is a worm and 100-tooth worm gear, the latter supporting the specimen to be measured. The worm is manually actuated by a knob which contains inside it three

Jan 1, 1961

By R. B. Booth, W. F. Linke

The continuous interest of the American Cyanamid Company in producing superior polymeric flocculants and dispersants for the mining industry has resulted in a broad, general study of the physical chemistry of mineral-polymer systems. The aim of this program has been to establish a sound theoretical basis on which practical problems could be analyzed, and which would lead to the development of new and improved flocculating and dispersing agents. This study has required the cooperative effort of physical, analytical, and polymer chemists, as well as engineers familiar with mining problems. This report summarizes the results of some of our investigations over the past two years. Further studies along these and other lines are in progress and will be reported in the near future. MEASURING FLOCCULATION IN THE LABORATORY The foremost problem in studying flocculation in the laboratory has been to establish reproducible methods of comparing and evaluating the effectiveness of various reagents. Previous workers have measured settling -rates 1,8 filtration rates2 , and sediment volumes3 . Direct observations of the floc size, or the clarity of the supernatant liquid, can also be used. AIthough these parameters are all interrelated, they do not always measure the same physical properties of the suspensions, and hence are not equally useful or reliable. Whatever the method of comparison selected. the testing procedure must be carefully defined and standardized. The results must be satisfactorily reproducible, and for practical purposes must accurately reflect the results obtained in mill use. In order to compare the data obtained in these various types of test, the physical nature of the flocculated system must be examined in some detail. In an unflocculated system, only the larger or denser particles (the "sands") settle at a useful rate. The sediment can be observed to build up from the bottom of the cylinder, while the fines remain in suspension for a long period of time. If the system is not disturbed, even the microscopic fines can be seen to classify themselves as they fall over a period of days. When a high molecular weight polymer is added to the system, the colloidal particles are tied together randomly in flocs, and the sands and fines settle together. The line of demarkation between solid and liquid phases moves downward from the top of the cylinder. The settling rate of the flocs depends on their size and on their effective density. For a given mineral sample, the larger the flocs, the faster the settling. Since the large flocs cannot pack as closely as small ones, they produce a sediment with more pore spaces, and hence of larger volume, which is easier to pump and to agitate. Although it is nearly impossible to measure individual floc sizes, it is relatively easy to determine settling rates and sediment volumes. Of these two, the settling rate is simpIer to measure, and hence generally more practical. Reproducible sediment volumes can only be obtained by allowing the solids to pack for a rather long period of time (e.g., hours to days), and although the volumes thus observed are directly related to the degree of flocculation (average floc size), it is often impractical to wait so long for the results.

Jan 1, 1961

By C. Elbaum

Specimens consisting of several crystals, each with a prescribed geometry and a controlled orientation with respect to both the extermally intposed stress and the highboring crystals, werer deformed in tension at room temperature. It is shown that the influence of a grain boundary on the plastic defolrmation of a mullicrystal depends on the extent to which the presence of the grain boundary, contributes to the necessity 01. multiple slip in the component crystals. MANY investigations were carried out in the past on the plastic deformation of aluminum single crystals, bicrystals, and polycrystals. The various studies of bicrystals aimed primarily at elucidating the influence of grain boundaries on the mechanisms of plastic deformation. In the earlier studies the emphasis was placed primarily on finding the "strength" contributed by a grain boundary. In more recent investigations it became apparent that the grain boundary does not contribute any strength per se but rather acts by influencing the mechanism of plastic deformation in the adjoining grains. The concept thus emerged of the plastic properties of a bi-crystal, and by extension of a polycrystal. depending solely on the mechanism of deformation prevailing in each grain. The contribution of the grain boundaries thus becomes limited to the type of restrictions imposed on each grain by its neighbors. The present investigation of the behavior of specimens consisting of more than two crystals was designed for the purpose of studying the influence of the grain boundary in simple aggregates. The findings of this study further support the concept that the restriction imposed on a deforming crystal by its neighbor constitutes the predominant factor in the plastic deformation of polycrystals. The term "multicrystal", in contrast with single and polycrystals, is used to describe specimens consisting of several crystals. Each one of the component crystals has a prescribed geometry and a controlled orientation with respect to both the externally imposed stress and the neighboring crystals. Two groups of multicrystals were prepared. All the component crystals of both groups of multicrystals had the axial orientation (isoaxial multicrystals) shown in Fig. 2, with the primary slip plane and slip direction both at 45 deg to the specimen axis. In the first group (hereafter referred to as group I) the component crystals were symmetrically oriented with respect to the plane of the grain boundary; the primary slip plane and slip direction were respectively at an angle of 60 and 30 deg to this plane. For the second group (group 11) the axial orientation remained the same and the relative positions of the primary slip plane and slip direction are shown in Fig. 3. EXPERIMENTAL METHODS The material used was aluminum 99.993 pct pure, according to the manufacturer. The specimens were produced by growth from the melt in a horizontal graphite container. In order to maintain the prescribed geometry of the individual crystals of a multicrystal, a technique described elsewhere was used.' The specimens were grown under a small, positive pressure of helium at a rate of approximately 1 mm per min. annealed in air for 24 hr at 640°C and cooled over a period of several hours to room temperature. All the multi-crystals were rectangular in cross-section, 6 by 18 mm (see Fig. 1) and from 150 to 170 mm long. In each group bicrystals, tricrystals and quadru-crystals were prepared as shown in Figs. 1 and 4.

Jan 1, 1961

Jan 1, 1961

By J. B. Cohen

EARLY investigators1-4 reported that additions of beryllium to silver caused embrittlement. The in-termetallic phases which are now known to exist5,' were not known at the time. Solidification appeared to occur by a simple eutectic reation.4,7 (The primary beryllium-rich phase appeared globular, when the liquidus temperature was near the melting point of pure beryllium, and angular when the liquidus was much lower; in discussion of Sloman's work,4 it was suggested that this difference was due to an intermetallic compound which was not revealed by the polishing technique.) In the investigation reported here, it was found that if the compounds now known to exist were avoided, certain beryllium-silver alloys containing large amounts of the primary beryllium phase possessed sufficient ductility to be cold formed. Alloys containing 14 and 20 at. pct Ag were prepared from vacuum-cast beryllium containing 1.0 to 1.2 pct impurities and silver, 99.99+ pct pure. The elements were melted together by induction in beryllia or alumina crucibles under argon at less than 1 atm pressure. The cooling rate was adjusted by regulating the temperature of the melt and by reducing the power to zero without stopping the water flow in the induction coils, which were in contact with the crucible. Ingots were about 0.75 in. in diam with a grain size of 0.12 to 0.25 in. Microscopic investigation of transverse or longitudinal sections of an ingot cast from a temperature well above its melting point showed that, in general, the compounds did not form, presumably because of rapid cooling; the alloys consisted of beryllium-rich particles in a silver-rich matrix as shown in Fig. 1. A eutectic structure was not observed, probably .due to the low beryllium concentration of the eutectic composition and eutectic divorcement. Within a given grain, the small globules of the primary phase were all oriented in the same direction, as indicated by examination under polarized light; the orientation varied from grain to grain. This suggested that the globules were branches of the same dendrites. If cooling was too slow, a third constituent was observed in the microstructure; because it formed around the primary beryllium-rich globules it was Fig. 1—Ag-Be alloy, 18.6 at. pct Ag, as-cast. Bright field illumination. Etched with an aqueous solution of NH,OH and H2O2. X500. Reduced approximately 21 pct for reproduction. probably the 6 compound. A thin ring around an ingot often contained this constituent. When this compound was present in large quantities, or the silver segregated so that large regions occurred with the beryllium-rich globules in contact, sections from the ingot were brittle. When the compound was initially absent, it was possible to carry out reductions in thickness of 50 pct by cold rolling with only limited edge cracking; failure did not occur until reductions of 80 pct. A typical cross section after cold rolling is shown in Fig. 2. After simple bend tests, microcracks were found; these apparently were difficult to propagate in the two-phase structure, as the specimens continued to behave in a ductile manner after the cracks appeared. Similar cracks would probably be closed during rolling. It is apparent in Fig. 2 however, that the primary beryllium particles undergo extensive plastic deformation. Hardness values of the alloys indicated that some work-hardening occurred during cold working. Fig. 2—Ag-Be alloy, 18.6 at. pct Ag, reduced 83 pct in thickness by cold rolling. Bright field illumination. Etched with an aqueous solution of NH,OH and H2O2. X500. Reduced approximately 21 pct for reproduction.

Jan 1, 1961

By V. D. Perry

Mineralized breccia pipes, because of their widespread occurrence and close structural relations to some of the world's great ore bodies, are objects of unusual interest for mining engineers and geologists. The literature contains many references to them, but it is questionable whether their genetic significance and economic importance have been sufficiently emphasized. The purpose here is to stress these features, relating them to the field facts, for the particular benefit of younger generations of geologists who, confronted with and sometimes confused by the growing flood of geochemical, geophysical, and other specialized research approaches, may be reassured that mappable field relations remain a foremost guide to a better understanding of ore deposits. A mineralized breccia pipe is a pre-mineral, breccia structure which has controlled the circulation and deposition of subsequently introduced mineralization. It is composed of relatively rotated angular or rounded rock fragments, set in a mineralized matrix. A pipe in plan outline may be circular, oval or approach polygonal form, with a steep to vertical axis proportionately much greater than its horizontal dimensions. The pipe is a steeply plunging, chimneylike mass of brecciated rock cemented with later minerals. Rock breaks in a variety of ways and complete fragmentation often occurs without rotation of individual pieces. A finely broken rock mass may fit into a tight jig-saw pattern, each fragment having mutually concordant boundaries with its neighbors. The result is a stockwork of innumerable reticulating cracks that, once cemented by mineralization, forms a complicated intersecting network of individually insignificant but collectively important seams and veinlets. Stockwork fracturing among its many forms takes the shape of domes of subsidence, fracture pipes, and related peripheral zones around and over breccia columns, or a combination of any of these structures. The significance of the mineralized breccia pipe is that it represents the extreme or climactic expression of a structural type which has a variety of mutations including subsidence domes, fracture pipes, and other stockwork zones all with related ancestry and similar to dissimilar characteristics. These allied and associated structures hold an answer to the fundamental question of the origin of many important ore deposits. CANANEA-TYPE LOCALITY FOR BRECCIA PIPES The Cananea district is characterized by an unusual development of mineralized breccia pipes. It is an important copper producer located in Sonora, Mexico, a short distance southwest of Bisbee, Arizona, and at the southerly limit of the great porphyry copper belt of the southwestern U.S. Cananea's rocks consist of Paleozoic quartzite and limestone capped unconformably by a thick series of volcanics including andesitic flows, tuffs, and agglomerates. These rocks have been intruded by a deep-seated granite with related basic and acid differentiates including dikes and plugs of quartz mon-zonite porphyry. Mineralization coincides with a northwesterly trending belt of these intrusives which break upward into and through the sedimentary-volcanic rock sequence. The district has weakly defined tectonic alignments in a northwesterly direction with subordinate intersecting fracture elements, but lacks important faulting or fissuring to provide throughgoing avenues for the upward circulation of mineralizing fluids. Thus, as will be discussed in subsequent paragraphs, the alternate way in which late magmatic and hydro-thermal derivatives of the parent magma reached the near-surface zone was by excavating their own breccia pipe channelways. There are numerous stages of breccia pipe development, related both in time and space to magmatic activity. A compilation of similarities and differences in various pipes suggests that proximity or remoteness of a demonstrable or inferred magmatic source provides an orderly genetic basis for describing the following representative types. Cananea Duluth Type: There are no intrusive rocks within or close to the Cananea Duluth pipe; therefore, the existence of any deep-seated magma that may have been related to its formation must be inferred. The structure is an oval-shaped ring 1200 x 300 ft in plan dimensions, cutting steeply across low angle, bedded tuffs, and other volcanics; it has been developed by drill holes to a depth of 2000 ft below the surface. The ore follows the periphery of the pipe and is composed of intensely brecciated rock which is cemented by minor galena, sphalerite, chalcopyrite, quartz, carbonates, and adularia. There is a definite vertical zoning of sulfides with less galena, continuing sphalerite, and increasing chalcopyrite at deeper levels. Within the interior of the ore ring, the brecciation becomes progressively weaker and coarser, the whole indicating relatively gentle slumping with broken, thin tuff beds pre-

Jan 1, 1961

Jan 1, 1961

By J. W. Goss, M. J. Coolbaugh

Most grouting has been done to stop water flaw in mines and for stabilizing foundations of various man-made structures, a survey of the U.S. literature reveals. Apparently Sun Manuel is one of the first mines in this country to use grout extensively underground for strengthening and stabilizing ground in drifts, shafts, and stations. A comparison with other procedures as well as details of the Sun Manuel program are covered. The employment of pressure grouting as a ground stabilizer at the San Manuel Copper Corp. mine reduces delays in both development and production, lowers costs, and makes possible safer working conditions. It specifically reduces delays in haulage operations and permits the maintenance of normal ventilation. In development work this grouting allows faster excavation by cementing together highly fractured or broken ground that otherwise would require extensive cribbing or spiling. In drift repair, it consolidates the loose or fractured rock over the timber or steel drift supports, thus decreasing the frequency of repair, lessening the hazards from falling rock, and curtailing delays due to blockage of drifts by muckpiles and repair operations. Pressure grouting is the process of pumping an accurately controlled mixture of cement and water into loose, fractured, or porous rock. The ratio of water to cement varies according to the nature of the rock encountered, from the thinnest mixture of 30 gal of water per sack of cement, used to fill very fine fissures, to very thick grout of 5 gal of water per sack of cement, used to fill large fissures or extensive areas of loose, broken rock. The pumping pressures at San Manuel vary from about 100 to 1000 psi, depending upon the compactness of the rock, while at other mines the pressures sometimes go as high as 5000 psi. In order to grout an area in which ground water is encountered, the pumping pressure of the grout must be increased by an amount equal to the pressure of the ground water. GROUTING IN DEVELOPMENT WORK The areas under development at San Manuel include many sections of loose, highly fractured rock. Prior to the advent of the grouting program, drifts, shafts, and stations in these areas could not be excavated without utilizing extensive support such as cribbing, spiling, and breast boarding. This slowed down the development work and increased the costs considerably over that required for excavating in more competent ground. Now, when a particularly bad area is contacted, the mining is temporarily stopped while the bad portion of drift or shaft is grouted. After the grouting is completed, the excavating is resumed with an approximate decrease of 50 pct in lost time and costs. Grouting in Drifts and Turnouts: The grouting procedures and patterns used in drifts and turnouts are very similar. The standard procedure for grouting such a turnout is shown in Fig. 1. After the ground is supported as well as possible with cribbing or lagging, the gaps between the back lagging and side lagging are plugged with empty cement sacks or additional timber, if necessary. The ground surrounding the proposed turnout is then grouted in two stages. In the first stage, the broken and caved rock is drilled and grouted to a depth of 5 to 10 ft, depending on the depth of the broken rock. This stage forms a grouted seal that allows higher pressures to be used at depths beyond 10 ft without developing excessive leaks at the face. After a grout hole is drilled, a 10-ft long pipe is wedged tightly into it with empty cement sacks. Grout is then pumped into the hole until the pressure reaches 200 to 300 psi, and the pipe is subsequently removed if it hasn't become cemented in. No more than one hole at a time is drilled and grouted because the grout has a tendency to go from one hole into another, plugging up the latter. Five or six holes are usually adequate for the first 10 ft of grouting. Experience has shown bentonite to be a useful admixture to the grout, particularly when it is indicated that the grout is being lost into large fissures or voids. Bentonite increases the plasticity of the grout enabling it to remain in place more easily until it has begun to set. In the second stage, the longer holes are drilled through the previously grouted rock, after which the

Jan 1, 1961

By C. E. Birchenall

DaVIES and Richardson1 have measured composition changes for Mn1-Owith variation in the equilibrium partial pressure of oxygen at 1500°, 1575°, and 1650°C, where 6 is the deviation from the simple stoichiometric ratio. No trend in this relationship with temperature was observed, so these remarks apply to the whole temperature range. The experimental values and their spread are shown by the bounded straight-line segments in the figure. Davies and Richardson assume that the stoichiometric oxide shows Frenkel disorder, given by the equation Mn2+Mn2+ + Mn" [1] where Mn2+ is a normal lattice ion which moves into an interstitial site (Mni2+) leaving a vacancy (Mn,"). The superscript dashes indicate that there is an excess charge on the anions surrounding the vacancy. Schottky disorder, in which a perfect crystal acquires equal numbers of cation and anion (Ov") vacancies according to leads to equations of exactly the same form in the analysis below. In view of the neutron diffraction results of Roth,2 which shows extensive Frenkel disorder in Fe1-o, Eq. [I] seems more appropriate. The deviations from stoichiometric composition, MnO, result from the addition of oxygen according to In effect Davies and Richardson ignore the variation of Mn3+in constructing equilibrium constants. This approximation might not appear to be very serious if the excess positive charges are really free posi- tive holes (P+) and if the intrinsic conductivity is high as a result of the reaction where e- is a conducting electron. However, the calculations of Buessem and Butler3 indicate that the partition functions of the free charges make a large contribution to the equilibrium constant, and this approximation would not be a desirable one. Furthermore, Morin4 shows that the positive holes are probably localized in MnO and therefore are described better by the mass action principle than by a degenerate gas model. Davies and Richardson's considerations require that activity corrections be applied beyond a 6 value of about 0.005. It is proposed to show here that the dissociation pressure curve can be calculated reasonably well over the whole composition range studied without activity corrections by choosing appropriate equilibrium constants for reactions 1 and 2. As stoichiometric composition The mass action constant, Ks, for reaction [I.] and the equilibrium constant, K,, for reaction [2] are given by where the concentrations which do not change significantly have been absorbed into the constants. The brackets denote concentrations. The deviation from stoichometric composition is given by gardless of [Mni2 ]. Substituting Eq. [7] in [6]

Jan 1, 1961

By J. C. Bokros

At temperatures above 950°F, zirconium which has been strained a critical amount will experience critical recrystalli-zation. The large grain size thus formed can result in a reduction in the fatigue life by a factor of 2 to a factor of 9 at both high and low temperatures. The critical strains for zirconium vary from 15 pct at 900°F to about 2 pct at 1125°F, while those for zircaloy III vary from 15 pct at 1050°F to 5 pct at 1200°F. Zircaloy II ,experiences no critical recrystallization up to 1200°F. ZIRCONIUM has become increasingly important for nuclear applications and is being employed in reactors at temperatures where recrystallization can occur. Recent work' on the fatigue properties of zirconium has shown that coarse-grained structures developed in zirconium during recrystallization have much lower fatigue lives than fine-grained material. The development of these large grain sizes under such conditions is well known to be the result of critical recrystallization.2-4 Briefly, critical recrystallization occurs when a metal is strained a critical amount and annealed in the recrystallization range. The critical strain is defined as that minimum strain necessary to promote nucleation in the time unit and temperature considered. The maximum grain size occurs at the critical strain, since at this strain the fewest stable nuclei per unit volume grow into new grains. Recrystallization is not evident below this; strain since the nucleation rate is vanishingly lolw, and metal strained more than the critical amount has a higher nucleation rate per unit volume which results in a finer grained structure. This paper describes the critical recrystallization observed in reactor-grade zirconium, zircaloy 11, and zircaloy I11 as a function of annealing temperature and temperature of strain. EXPERIMENTAL The zirconium used was hafnium-free, double arc-melted sponge obtained from the Bureau of Mines. The ingots were forged and hot reduced to 0.08 to 0.10-in. sheet and subsequently cold reduced to 0.035-in. sheet with intermediate and final anneals at 1200 OF for 15 min. The zircaloy II and zircaloy III used were 0.035-in. sheets in the annealed condition. Compositions of these materials as tested are given in Table I. To determine the critical strain at a given temperature, tensile specimens having a0.065-in. taper in a 2-in. gage length were machined from the sheets parallel to the rolling direction. The specimens were polished, etched, and examined metallographically to ensure that all of them were strain-free, initially. Tukon indentations were made every half millimeter along the gage length of each specimen. These served as fiducial marks enabling the plastic strain to be measured with a microscope and micrometer stage. The specimens were pulled to failure at a rate of 0.002 in. per min in an Instron tensile-testing machine. Three temperatures were used for straining: room temperature, 500 ºF, and 900 ºF. The 500º and 900 ºF tests were run in an argon atmosphere. Since the rate of work hardening changes with temperature, a different critical strain was obtained for each temperature of strain. Specimens, prepared in the manner described, were annealed in vacuunl at various temperatures ranging from 900" to 1200 F for up to 1000 hr, after which they were examined metallographically. The critical strain and the recrystallized grain size as a function of strain were determined for each specimen. Results for the same temperature and time indicated that the critical strain measured was reproducible to within 5 pct of each other. RESULTS AND DISCUSSION Since increasing the time of anneal increases the probability of the nucleation of a new grain, the critical strain decreases with time at a decreasing

Jan 1, 1961

[Appalachian Petroleum Section: James M. Kimbrough, chairman; John H. Adair, first vice-chairman; Walter L. Diddle, second vice-chairman; Byron M. Finch, secretary-treasurer. Balcones Section: Robert V. West, chairman; Frank W. Jessen, first vice-chairman; Zane V. Morgan, second vice-chairman; Brantley G. Myers, secretary-treasurer. Meets 4th Wed., Petroleum Club (San Antonio) or Terrace Dining Room (Austin) Billings Petroleum Section: Charles B. Evans, chairman; L. M. Lyon, first vice-chairman; Gordon E. Lansford, second vice-chairman; Fred. A. Haddenhorst, secretary-treasurer. Meets 3rd Tues., Billings Petroleum Club. California Coastal Section: Robert L. Richardson, chairman; David K. Hayward, first vice-chairman; W. Jackson Bransford, secretary; William N. Harwell, treasurer. Meets 3rd Thurs., Pierport Inn, Ventura. Caracas Petroleum Section: Edwin A. Moo re, chairman;. Jose Cirigliano, vice-chairman; Frank B. Davis, secretary; Robert N. Powers, treasurer. Meets 2nd Tues., Caracas Theater Club.]

Jan 1, 1961

By H. L. Hartman

The estimation of the magnitude of pressure changes which occur in mine ventilation circuits is of primary importance to the mining engineer in making changes in an existing mine or in projecting the ventilation requirements of a new mine. It is a formidable task in many cases, because no adequate theory is available and empirical data are unreliable or even lacking. The common practice of increasing one's estimate of the mine static pressure by an arbitrary "safety factor" to compensate for the unknown is a poor substitute for accuracy. Generally little difficulty is encountered in calculating friction losses in mine airways. As long as the friction factor can be reliably estimated, the well-known Atkinson formula gives accurate results. However, it is in the determination of shock losses that the ventilation engineer encounters difficulty and is apt to resort to an arbitrary allowance. Shock losses may constitute as little as five or as much as twenty-five percent or more of the overall pressure drop in the mine, and such allowances are generally completely unsatisfactory. Since other shock losses have been covered by McElroy (1), this paper deals with a long-neglected source of shock and pressure change in ventilation circuits, divergence and convergence of airflows, here referred to by their more common mining terms, splits and junctions. PREVIOUS WORK There is practically no fundamental information available in ventilation literature on the splitting and junction of airflows, either in mining or mechanical engineering periodicals or handbooks. Fluid mechanics textbooks seemingly avoid the subject of divergence and convergence, except for brief empirical treatment. The only investigation reported for mine ventilation systems was by Weeks, et a1 (21), in 1933. Although applicable to the case of line brattices in coal mines, it cannot be applied to the general case of mine splitting and junctions involving change of direction with little or no area change. In dealing briefly with the subject of shock losses at splits and junctions, McElroy (1) suggests only that they be considered comparable to bends, with losses at junctions being increased 50 pct to allow for interference of merging streams. He recognizes the importance of quantity ratio but concludes that exact quantitative data are lacking to compute losses at splits and junctions. A very capable summary and analysis of the state of knowledge in industrial ventilation and fluid mechanics regarding losses due to divergence has recently been reported by Gilman (3). In considering all earlier work dealing with both air and water as fluids, he compares results on a common basis and obtains good empirical agreement. For divergent flow, the controlling variable in each branch is the quantity ratio, although the deflection angle is of secondary importance. Empirical equations are presented for straight and 90-deg branches with varying velocity ratios, assuming branches smaller than the main duct. These formulas are modifications of the Borda equation for abrupt contraction and expansion. The only information available on shock losses in convergent flow are approximate experimental data presented by Alden (4) and the Manual of Recommended Practice for Industrial Ventilation of the American Conference of Governmental Industrial Hygienists. Alden considers the effects of both quantity ratio and deflection angle. The data indicate an effect opposite to that observed in divergence, the shock loss varying directly with quantity ratio. SHOCK LOSS THEORY Because pressure losses due to shock have been found to bear a constant relation to the mean velocity of flow in a given conduit, they are frequently expressed as a dimensionless function of the velocity head, termed the shock loss factor. For airflow in ducts and mine airways, the shock loss may be represented by the formula (1): Hx - XHv, (1)

Jan 1, 1961

By J. W. Cunningham, W. A. Tiller, D. H. Lane

The effect of ultrasonic vibrations on ingot solidification has been considered both theoretically and experimentally. The theoretical section elucidates the mechanisms by which the ultrasonic vibrations might refine the ingot structure. The experimental section describes a very efficient means of introducing the ultrasonic vibrations to the freezing interface. The application of this technique to the consumable-electrode arc melting process shows that the columnar mode of freezing may be effectively suppressed yielding an ingot structure consisting predominantly of equiaxed grains. The results support the contention that the ultrasonic vibrations increase the nucleation frequency in the layer of liquid adjacent to the freezing interface. CERTAIN structural characteristics of as-cast ingots are undesirable since they create considerable difficulty for subsequent processing and often produce defects in the finished product. These structural defects are illustrated in Fig. 1(a), and are: 1) planes of weakness where columnar grains growing from the different mold faces meet, 2) strength anisotropy due to the preferred orientation of the columnar grains and 3) inhomogeneity due to segregation at dendrite boundaries within the columnar grains and at the columnar grain boundaries. The type of ingot structure that eliminates 1) and 2) and diminishes 3) is the fine-grained equiaxed structure illustrated in Fig. 1(c). Many attempts have been made to produce the desirable ingot structure illustrated in Fig. 1(c) by inoculating, casting with varying superheats, vibrating or stirring, or by irradiation with ultrasonic vibrations. These attempts have met with various degrees of success, the ingots generally having the structure illustrated in Fig. 1(b). All four methods have been the subject of numerous investigations,1"4 the latter producing the most erratic behavior and inconsistent success. The method of ultrasonic irradiation during ingot solidification has shown considerable promise on small-scale ingots (diam 2 in.) but has not yet been successfully adapted to large-scale operation. The present paper considers the application of ultrasonic radiation to ingot solidification both theoretically and experimentally. The purpose of the paper is to elucidate the mechanism by which ultrasonic energy refines the ingot structure and to describe a method for producing the desirable ingot structure in both laboratory scale experiments and in ingots of commercial size. THEORY Conventional Casting—Before discussing the possible effects of ultrasonic radiation upon ingot structure let us first briefly review the reasons for the breakdown of columnar crystallization to equiaxed crystallization during ingot solidification. The following is a slightly oversimplified description. In the conventional method of casting, the alloy having a certain degree of superheat, T, is poured to fill a cold mold which is at room temperature, Tc,. Due to conduction of heat through the mold wall the outer rim of liquid metal cools rapidly to the liquidus temperature and begins to supercool. At some degree of supercooling, nuclei of solid form at random in this supercooled layer at the mold surface. The grains in this chill layer begin to grow inwards producing a solute build-up, thus lowering the freezing temperature at the interface as illustrated in Fig. 2. As growth proceeds, the temperature gradient in the solid conducts away not only the latent heat of fusion evolved during growth but it also conducts away some of the superheat of the melt so that the temperature

Jan 1, 1961

By R. W. Stahl

Employees and dollars are necessary to all enterprises and any force, such as fire, which destroys either, can bring very serious consequences, including business failure. Since everyone acknowledges the necessity of conserving human life, we will devote most of our paper to the economic factors involved. A coal-mine fire is considerably more difficult to combat than a fire in a surface building because of the combustible surroundings in the mine, the explosive gases emitted from the coal, the suspension of explosive dust when mine timbers burn and the unsupported roof falls, the limited space in which approach to the fire can be made, and the limited avenues of escape. A survey of coal-mine-fire protection indicated that a rather low value was being placed on property and equipment. This prompted comparison with other industries. The comments made in this comparison are not intended as an indictment of the coal industry, since we do have mines that are well protected against fire, but the comments on conditions at the mines visited are factual and may help others to improve similar conditions. SURVEY PROCEDURES Plants of the steel, aluminum, rubber, chemical, and petroleum industries were surveyed, and all facets of the fire-protection scheme were covered. The industrial plant invariably had four times as much water readily available for instant use than the mine. In many mines where water is available, no equipment for applying it is provided, such as hose, nozzles, and adapters, so that time is lost in collecting this equipment. (See Table I.) EXTINGUISHERS Many more and generally larger portable extinpishers are available at the plant as compared to the mine. The extinguishers are generally placed nearer to sources of danger. In addition, respiratory protective equipment is always available at strategic points in plants but only in a few mines. (See Figs. 1 and Table 11) INSPECTING AND TESTING EQUIPMENT Plant water facilities are checked weekly and flushed at 6-month intervals, and hose is pressure-tested at least yearly; some mines check water yearly, but hose is not tested unless used. Fire extinguishers are inspected weekly in plants and completely overhauled yearly; one mine had monthly checks and overhauls only after use. FIRE BRIGADES AND TRAINING Every plant had designated fire-fighters and some complete full-time brigades; at one mine a chief mechanic was designated as being in charge of fires, and other depended entirely upon officials who might not be familiar with the equipment. COST OF FIRES AND PERCENTAGE OF CAPITAL INVESTMENT IN FIRE PROTECTION The direct cost of recent mine fires ranged from $128,000 to $240,000; indirect costs, such as lost tonnage, idled equipment, lost orders, and the like might run the cost to over $500,000. (See Figs. 2 and 3).

Jan 1, 1961

By J. E. House, C. J. Lewis

In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-pregnant sulfuric acid leach solution is contacted with an organic extractant dispersed in kerosene; the uranium enters the organic phase and associates itself with the organic extractants. Simultaneously an-ions or cations, as the case may be, leave the organic extractant phase and enter the aqueous phase. The two phases are then allowed to separate by virtue of their immiscibility. Based on the data available, a generalized procedure for the recovery of molybdenum from the sodium carbonate scrub liquor of acid leach uranium mills using secondary and tertiary amine extractants is presented, as is another generalized procedure for the recovery of molybdenum from the bleed stream of alkaline leach uranium mill circuits. Such economic appraisal as can be made at this stage of the development is also indicated. Molybdenum occurs in most uranium ores. Whether the uranium is solubilized by the acid leach route or by the alkaline leach process, some of the molybdenum present in the ore is solubilized along with the uranium. Molybdenum in uranium mill pregant solutions may cause processing difficulties, among which may be mentioned the poisoning of resins, the precipitation of organo-molybdenum complexes, or the build-up of molybdenum in recycle streams. While the uranium milling industry has been successful in applying measures to overcome the problems arising from the presence of molybdenum in the circuits, such measures nevertheless add to the cost of producing uranium yellow cake concentrate. Table I shows the analysis of typical uranium pregnant leach solutions. Uranium recovery by solvent extraction is presently practiced only in conjunction with acid leach circuits, and it may be noted that of about 16,000 tons of uranium ore daily processed by sulfuric acid leach, more than half of this tonnage follows the solvent extraction route. Uranium mills leaching ore with sodium carbonate-sodium bicarbonate solution (the so-called alkaline leach circuit) do not use solvent extraction. In this case the uranium pregnant solution is sufficiently pure to permit the direct precipitation of uranium concentrate with subsequent recycle of the alkaline leach solution. While solvent extraction procedures may be developed which are economically attractive for the alkaline leach uranium process, this has so far not been the case. The earlier uranium mill solvent extraction circuits used organo-phosphate extraction reagents, among which di-2-ethylhexyl phosphoric acid was popular. It should be borne in mind that uranium in its sulfuric acid leach solution exists in an equilibrium involving uranium in both anion and cation forms: uo + 2S0 = [uo2(so4)2] For this reason uranium may be extracted by either cation extractors or anion extractors and this is likewise true of molybdenum. However, the organo-phosphate extractants also load ferric iron which is usually present in the uranium acid leach solution. For this reason, it is necessary to reduce this solution substantially to the ferrous state ahead of solvent extraction using organo-phosphate reagents. The amines, being anion extractors, selectively reject iron and thus render the iron reduction step unnecessary. In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-

Jan 1, 1961

By K. E. Merklin, F. D. DeVaney

Students of the modern blast furnace seem unanimously agreed that they are observing a major revolution in practice. Rather than changing construction and operation of the furnaces, most of the great advances now under way deal with the raw materials feed. Prepared burdens have gained increasing significance during recent years. Sizing, concentration, and agglomeration have all become bywords to greater production and reduced operating costs. The remarkable results with test furnaces in which self-fluxing sinter makes up part or all of the ore charge foretell even greater improvements. Other investigators have reported many of these advances, and it is enough to predict here that more and more operators will turn to the use of self-fluxing sinter. This change in the operator's requirements comes at a time when one major section of the iron ore regions, the Mesabi, is finding it difficult to maintain, let alone improve, the tonnage and grade of its shipping product. To meet the new demands with respect to quality and structure, the iron ore producers of this area turned first to screening, crushing, and concentrating intermediate ores. More recently, declining higher-grade reserves have led to multimillion-dollar investments in plants to treat the low-grade magnetic taconites.1,2 To yield a suitable concentrate, the Minnesota magnetic taconites must be ground to about 80 pct through 325 mesh, too fine to be charged directly to the blast furnace or to make acceptable feed to the sinter strand. The pelletizing process was developed to meet the structure requirements laid down by the blast furnace operators. Taconite producers are succeeding in their efforts just when the attention of the blast furnace man is being drawn to the great advantages of self-fluxing sinter.3 The taconite operators are ready with an alternative. Why not produce sclf-flus ing pellets? At the Hibbing laboratory of Pickands Mather & Co., where much of the early work was done that led to Erie Mining's 7.5-million ton taconite plant at Hoyt Lakes, Minn., preliminary investigations have been made. Present Taconite Operations: Details of this work will be clarified by a brief review of agglomeration practice in commercial taconite plants. After the taconite has been ground to about 80 pct through 325 mesh to liberate the magnetite grains, the ore is concentrated on magnetic separators, thickened, and filtered to about 10 pct moisture. This structure and moisture fit in closely with the conditions required to produce good balls. It has been found with taconite concentrates that at least 60 pct of the feed to the balling drum must be —325 mesh to obtain good ball compaction within reasonable time. The filter cake moisture usually permits a slight water addition at the balling drum, and further balling control is obtained by adding about 12 Ib of bentonite and 1 1/2 lb of soda ash per dry short ton of concentrate. These additives, and occasionally small quantities of fine coal, are thoroughly mixed with the filter cake ahead of balling. The swelling property of bentonite improves the green and dry strength of the balls so they can be handled without breakage until they are fired. Soda ash is added for pH control to insure maximum performance of the bentonite, and coal is added for more heat in the firing process when required. The balls, normally 3/8 to l-in. diam, are brought to a temperature of 2350" to 2500°F in the indurating furnace. Heat for this process comes from three sources: 1) fuel burned in the combustion chamber, 2) oxidation of the magnetite concentrates to hematite, and 3) coal added to the balls. Pellet strength is developed by grain growth and bridging and, if enough slag-forming constituents are present, by slag bonding. Conversion to hematite is usually close to 100 pct in the fired pellets. Batch Test Procedures: Initial studies of self-fluxing pellets were made by producing balls on a batch basis under carefully controlled laboratory conditions. The l-in. balls, uniform in size, were fired in an oxidizing atmosphere in a standard laboratory muffle furnace. The firing procedure consisted of

Jan 1, 1961

By G. E. Keller, G. D. Coe

Several methods have been devised for rapid determination of sulfur in coal using a high-temperature combustion furnace. The fundamental principles of the various methods are similar but the techniques vary widely. The method described herein is no different in principle from some of its predecessors but varies considerably in technique. Originally devised to overcome the low-accuracy criticism leveled against the others, the tube-furnace method has proved successful for limited applications. Work is being done to widen the applicability of the technique. Subsequent to the work of A. Vita1 in 1920, a number of investigators have published data dealing with the rapid determination of sulfur in coal using a high-temperature combustion furnace.2-5 These, together with several unpublished methods, involve wide differences in technique although they are all similar in principle. Investigation of typical details of procedure showed that poor precision or accuracy could be attributed to one or more of the following: 1) inadequate quantity of sample to be representative, 2) a delivery tube temperature too low to confine condensation to a washable area of the tube, 3) no provision for back washing the delivery tube, 4) incomplete absorption of the sulfur compounds due to inadequate bubbler, and 5) the false assumption that all of the sulfur is evolved as sulfur dixoide. The method herein discussed is no different in principle from some of its predecessors; the differences are in technique. Although the method is rapid—a determination can be completed in 5 min— accuracy has riot been sacrificed in order to achieve speed. MATERIALS AND EXPERIMENTAL WORK Fig. 1 shows the arrangement of the equipment. A metered volume of oxygen is passed through calcium chloride for removal of moisture, then through Ascarite for removal of carbon dioxide, and thence through a glass tee to the feed end of the combustion tube. Connected to the side arm of the glass tee is a rubber tube that is closed with a pinch clamp during combustion of a sample. The products of combustion are expelled through a Vycor bubbler, the end of which is submerged 3 cm below the surface of a 1 pct hydrogen peroxide solution contained in a 300 ml high-form beaker. At the completion of the 4-min combustion period the contents of the beaker are titrated to a yellow end point, using a methyl red indicator, with 0.05 N sodium hydroxide (or sodium tetraborate). The pinch clamp on the tube leading from the tee in the oxygen line is then opened and mouth suction applied to draw the hydrogen peroxide up into the bubbler past the ell, almost to the rubber tube connecting the bubbler to the combustion tube. This washing of the bubbler is repeated until no further acid reaction with the indicator is obtained. The hydrogen peroxide-sul-furic acid solution is then titrated to the final yellow end point. Test work conducted at the Robena Laboratory disclosed the following critical conditions that had important bearing on the precision of the method: Weight of Sample: Several series of tests wherein the sample weight ranged from 0.1 to 0.5 g of -60 mesh coal showed that the lowest variance (standard deviation squared) resulted from the use of 0.5-g samples; the variance was slightly greater when 0.3-g samples were used, and with less than a 0.3-g sample weight the variance increased rapidly with decrease in weight. Although the use of a 0.5-g sample invariably resulted in a very slight explosion as soon as the sample entered the hot zone of the combustion tube, this did not detract from the precision of the method. Temperatures: Mott and Wilkinson5 found that a combination of low temperature (1250°C) and low oxygen flow rate (less than 700 ml per min) tended to give high results due to formation of oxides of nitrogen. Although the method herein employs a much higher oxygen rate than 700 ml per min, as an additional precaution to assure accurate results, 1315°C (2400° F) is recommended as the most desirable furnace temperature. The temperature of the elbow of the bubbler should be at least 75°C to minimize condensation

Jan 1, 1961