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By G. C. Broadhead

The lead district of Southeast Missouri covers an area of over 3000 square miles, including Maries County on the west, Jefferson on the east, Franklin on the north, and part of Madison on the south, or parts of ten counties. A general section of the rocks of the southeast part of this region would be about as follows, numbering from the top: 1. Sandstone (the second of Missouri geologists),.... 20 feet. 2. Chert beds—beds of passage below,...... 126 " 3. Magnesian limestone, chert, and quartzite,. .. 100 to 300 " 4. Lower magnesian limestone,......100 to 150 " 5. Gritstone and lingula beds,....... 50 " 6. Ozark marble beds,........5 to 20 " 7. Sandstone and conglomerate,. .. .. 5 to 90 " 8. Porphyry, Archæan 9. Granite, The thickness is approximate, for it is not possible to obtain the correct thickness of the various beds. No. 1 may be seen as detached outliers or loose masses of a hard sandstone or quartzite, and is found on the hilltops in the southern part of Madison County, in like topographical position in Reynolds County and the western part of Washington County. No. 2 consists of alternations of chert, clay, and quartzite. It is the formation which contains most of the limonite deposits of Central and Southern Missouri. In Reynolds County shafts have penetrated it 75 feet. Outcrops there and in Madison would indicate it to be at least 125 feet thick. Fossils obtained in various places prove it clearly to be of the age of the calciferous saudrock of the New York system, and Nos. 1 and 2 may also be referred to the age of the second sandstone of Missouri. No. 3 consists chiefly of thick beds of ash, drab and flesh-colored magnesian limestone, both coarse and fine-grained. These beds, where they occur, in Washington, Jefferson, and the southern part of Madison Counties, contain a good deal of quartz in drusy cavities. In Washington County the limestones are beautifully ramified by a system of connected drusy cavities generally lined with minute quartz crystals arranged in botryoidal and mammillary forms, and called " mineral blossom " by the miners. This is the lead-bearing rock of Washington County, but in the southern part of Madison it does not appear to be galeniferous. It is also undoubtedly the equivalent of

By G. C. Broadhead

THE lead district of Southeast Missouri covers an area of over 3000 square miles, including Maries County on the west, Jefferson on the east, Franklin on the north, and part of Madison on the south, or parts of ten counties. A general section of the rocks of the southeast part of this region would be about as follows, numbering from the top 1. Sandstone (the second of Missouri geologists), - 20 feet. 2. Cbert beds-beds of passage below, - 125 feet 3. Magnesian limestone, chert, and quartzite, - 100 to 300 feet 4. Lower magnesian limestone, - 100 to 150 feet 5. Gritstone and lingula beds, - 50 feet 6. Ozark marble beds, - 5 to 20 feet 7. Sandstone and conglomerate, - 5 to 90 feet 8. Porphyry, Archaean. 9. Granite, The thickness is approximate, for it is not possible to obtain the correct thickness of the various beds. No. 1 may be seen as de¬tached outliers or loose masses of a hard sandstone or quartzite, and is found on the hilltops in the southern part of Madison County, in like topographical position in Reynolds County and the western part of Washington County. No. 2 consists of alternations of chert, clay, and quartzite. It is the formation which contains most of the limonite deposits of Central and Southern Missouri. In Reynolds County shafts have pene¬trated it 75 feet. Outcrops there and in Madison would indicate it to be at least 125 feet thick. Fossils obtained in various places prove it clearly to be of the age of the calciferous sandrock of the New York system, and Nos. 1 and 2 may also be referred to the age of the second sandstone of Missouri. No. 3 consists chiefly of thick beds of ash, drab and flesh-colored magnesian limestone, both coarse and fine-grained. These beds, where they occur, in Washington, Jefferson, and the southern part of Madison Counties, contain a good deal of quartz in drusy cavities. In Washington County the limestones are beautifully ramified by a system of connected drusy cavities generally lined with minute quartz crystals arranged in botryoidal and mammillary forms, and called u mineral blossom " by the miners. This is the lead-bearing rock of Washington County, but in the southern part of Madison it does not appear to be galeniferous. It is also undoubtedly the equivalent of

Jan 1, 1877

POLITICAL developments, the impact of Korean warfare necessitating immediate rearmament plans, and recent developments in mining taconites have focused attention on their immediate exploitation to fill the gap created by diminishing reserves. The question of rich foreign ores versus low grade development is past-we need both sources with emphasis on the immediate large scale mining of taconites. In the past three months several things have occurred which add to the taconite picture: (1) Pellets have been used successfully in the blast furnace by Armco. (2) In Congress HR 8512 was introduced by Congressman Blatnik of Minnesota to amend the lending powers of RFC to permit long term loans on basic mining developments, particularly the taconites, which may be needed for the strengthening of the national economy. (3) Walter S. Tower, president of the American Iron & Steel Institute, announced that American steel companies are planning expansion of 6,000,000 tons of steelmaking capacities in the next few years. (4) The outbreak of war in Korea has obliged the United States to rearm and make plans for a semipermanent defense establishment to which 10 to 12 pct of the national income will be budgeted. (5) Startling new advances have been made in the Linde Air Products Co.'s jet-piercing drilling. The United States is now confronted with the problem of increasing its steel capacity and at the same time of securing enough iron ore to supply this increase as well as make provision for the rapidly decreasing supply of rich iron ore on the Mesabi range. Taconites, the low grade iron ore at the east end of the Mesabi, of which there are literally billions and billions of tons of easily concentratable magnetic ore, are the immediate answer. According to the Wall Street Journal on July 24: "Reserve Mining Co., backed by such steelmakers as the Armco Steel Corp., Wheeling Steel Co., Republic Steel Corp., is reported ready-as soon as the pending bill clears (HR 8512)-to seek a $100 million federal loan to build a taconite processing plant at Beaver Bay, Minn., to produce about five million tons (of pellets) annually of 60 pct to 65 pct in iron ore . . ." Experimental work indicates that pellets may be the final form, rather than nodules, agglomerates, or sinter. At Ashland, Ky., the Reserve Mining Co., which has obtained a

Jan 9, 1950

By D. K. Das, P. A. Beck, S. P. Rideout

IN a previous publication1 1200°C isothermal phase diagram sections were given for the Cr-CO-Ni, Cr-Co-Fe, Cr-Co-Mo, and Cr-Ni-Mo ternary systems, in which the a phase formed narrow, elongated solid solution fields. The present investigation is concerned with the 1200°C isothermal sections of the Co-Ni-Mo, Co-Fe-Mo, and Ni-Fe-Mo ternary systems. A prominent feature of these systems is the presence of narrow, elongated µ phase fields. The crystal structure of the phase designated as µ both here and in the previous publication1 was determined by Arnfelt and Westgren.2 For the (CO, W)µ phase, named by them Co,W, (and also frequently designated as a), these authors found that the crystal system is hexagonal-rhombohedra1 and the space group is D53d — R3,. Westgren and Mag-neli3 later found that isomorphous phases exist in the Fe-W and the Fe-Mo systems (these phases are often referred to as < and E, respectively). Henglein and Kohsok4 stated that the phase described by them as Co7Mo,; (otherwise frequently designated as c) is also isomorphous with the above three. The Co-Fe-Mo system was investigated at 1300°C by Koester and Tonn,5 who found a continuous series of solid solutions between (Co, MO)µ and (Fe, MO)µ Koester6 also indicated similar uninterrupted solid solutions in the Ni-Fe-Mo system. However, since the Ni-Mo binary system does not have a phase isomorphous with F, Koester&apos;s diagram is expected to be erroneous. No data appear to be available in the literature concerning the Co-Ni-Mo system. The face-centered cubic (austenitic) solid solut,ions of iron, nickel, and cobalt, which are quite extensive in all three systems at 1200°C, are here designated as the a phase. The body-centered cubic (ferritic) solid solutions, based on iron, are designated in this report as the ? phase, in conformity with the nomenclature used previously.&apos; Experimental Procedure The alloys were prepared by vacuum induction melting in zirconia and alumina crucibles. The lot analyses for the metals used have been given.&apos; The number of alloys prepared was 46 for the Co-Ni-Mo system, 65 for the Co-Fe-Mo system, and 113 for the Ni-Fe-Mo system. The compositions of these alloys were selected with due regard to maximum usefulness in locating phase boundaries. The alloy specimens were annealed at 1200°C in an atmosphere of purified 92 pct helium and 8 pct hydrogen mixture. Alloys consisting almost entirely of the face-centered cubic austenitic a phase, or of the body-centered cubic ferritic c phase were double-forged with intermediate annealing. The double-forged specimens were then final annealed for 90 hr at 1200 °C and quenched in cold water. Alloys containing considerable amounts of any of the other phases could not be forged. Such specimens were annealed for 150 hr at 1200°C and quenched. Microscopic specimens of all alloys were prepared by mechanical polishing, in many cases followed by electrolytic polishing. Description of the polishing and etching procedures used and tabulation of the intended compositions of the alloys prepared are being published in two N.A.C.A. Technical Notes.7,8 , Many of the alloys were analyzed chemically and, in general, the results are in excellent agreement with the intended compositions. X-ray diffraction samples were prepared by filing or crushing homogenized alloy specimens and by reannealing the obtained powders in evacuated and sealed quartz tubes. After annealing for 30 min at 1200°C the tubes were quenched into cold water. X-ray diffraction patterns were made with unfiltered chromium radiation at 30 kv, using an asymmetrical focusing camera of high dispersion. X-ray diffraction and microscopic methods were used jointly to identify the phases present in each specimen. The amounts of the phases in each alloy were estimated microscopically. The phase boundaries were located by the disappearing phase method. The results were used to construct 1200°C isothermal sections for the three ternary phase diagrams. The accuracy of the location of the phase boundaries determined in this manner is estimated to approximately ±1 pct of each component. The portion of the three phase diagrams lying between the µ, P, and 6 phases on the one hand, and the molybdenum corner on the other, has not been investigated. Recently Metcalfe reported0 a high temperature allotropic form of cobalt on the basis of dilatometric results and of cooling curves. In the present work no attempt was made to search for the new phase in the cobalt corner of the Co-Fe-Mo and Co-Ni-Mo systems. No alloy was prepared with more than 80 pct Co; the alloys used were intended to locate the boundary of the a phase saturated with cL. The microstructures of the quenched a alloys near the cobalt COrner gave no suggestion of an in-suppressible transformation On quenching. The location of the boundaries of the a + ? two-phase fields in the Fe-Ni-Mo and Fe-CO-MO systems was determined entirely by the microscopic method. The face-centered cubic a alloys near the ? field transform partially or wholly into the body-centered cubic ? phase on quenching from 1200°C to room temperature. The ? formed in this manner has an

Jan 1, 1953

By C. T. Butler, W. W. Kay, R. D. Boddorff, R. L Ash

DRILLING and blasting in anthracite open-pit mines is a continuous problem to contractors and explosive engineers because of the diverse conditions caused by the nature of the geological formations, the extensive mining of the portions of coal beds near the surface, and the proximity of many strip pits to populated areas. Pennsylvania anthracite occurs in four separate long and narrow fields totaling only 480 sq miles. The coal measures are rock strata and coal beds that are considerably folded and faulted. The crests of the anticlines are eroded extensively. The beds outcrop on the mountain sides and dip under the valleys. At first only the upper portions of the syn-clines could be stripped. Now stripping to increasingly greater depths is economically possible, as is indicated by the fact that the proportion of freshly mined anthracite produced by strip mining has increased from 3.7 pct of the total tonnage in 1930 to 29.6 pct in 1950. Much of the rock overlying the deeper beds now being stripped is so extensively broken that considerable difficulty is experienced in drilling satisfactory blast holes and in using explosives in such manner as to insure a uniformly broken material easily removed by the excavating machinery. Such breaking of rock strata has occurred because the bed now being stripped has been mined extensively in former years by underground methods, and tops of gangways and chambers have subsequently failed. Draglines are used to uncover coal where the overburden can be moved with little or no re-handling. These machines range in size from those having a 2 cu yd capacity bucket on a 60-ft boom to those handling a 25 cu yd bucket on a 200-ft boom. Draglines are also used to strip to the bottom of the coal basins if the depth and the distance between the crops are not too great. For this type of operation blast holes are drilled full depth to the bed. These holes are commonly 30 to 90 ft deep; however, in exceptional cases, holes may be as shallow as 12 ft or as deep as 130 ft. Drilling is normally done for blasts of 12,000 to 60,000 cu yd of overburden, 30,000 cu yd being considered an average blast if vibration is not the controlling factor. Where the stripping of wide basins or the exposure of a moderately pitching vein makes the use of draglines impractical, dipper front shovels equipped with 4 to 6 cu yd buckets load into trucks. Overburden is removed in benches of 25 to 30 ft with blast holes drilled 4 or 5 ft deeper than the planned floor of the bench. For shovels under 5 cu yd bucket capacity the volume blasted varies from 8000 to 12,000 cu yd, whereas a volume of 30,000 to 50,000 cu yd of overburden is frequently blasted at one time for the larger shovels where vibration is not an important factor. During the past decade the churn drill, generally the Model 42-T Bucyrus-Erie blast hole drill equipped for drilling 9-in. diam holes, has become the most common blast hole drilling machine. Electricity powers half the churn drills in use and is preferred on the large strippings where electric shovels are operated and the working area is concentrated. On these operations the cost of additional electricity for the drills is less than the cost of fuel to operate diesel units because of the existing large demand load of the excavating equipment. Moreover, electric motors start more easily in cold weather and generally are less expensive to maintain. Diesel driven units are employed where a higher degree of mobility is required. The average drilling speed is 8 ft per hr, although in softer rocks a rate of 15 ft per hr is attained. Where rock is hard and strata is badly broken, drill speeds may be less than 2 ft per hr. Low drilling production results under these circumstances when loose material falling from the upper portion of the drill holes causes drill stems to be jammed. Rock formations vary so greatly in the region that a 9-in. diam churn drill bit may become dull after drilling only 2 ft or may drill satisfactorily for 56 ft; however, an average of 35 ft is usual in sandstone of medium hardness. Dull bits are hoisted to flat bed trucks by the sand line of the drill and are usually sharpened in the contractor&apos;s bit shop adjacent to the job. Care is generally taken to cover the thread end of the bit with a cap. To facilitate handling of bits around the drill, a heavy thread protector having an eye top is becoming more popular than the flat-top rubber or metal cap furnished with new bits. The 9-in. diam blast holes for a 25 to 30 ft bench are normally on 18x18 ft to 20x20 ft spacings, depending on the character of the overburden, although in broken ground 15x18 ft centers may be used to obtain better breakage and a more even bottom for the bench. The patterns of holes for shots

Jan 1, 1953

By R. D. Boddorff, R. L. Ash, C. T. Butler, W. W. Kay

DRILLING and blasting in anthracite open-pit mines is a continuous problem to contractors and explosive engineers because of the diverse conditions caused by the nature of the geological formations, the extensive mining of the portions of coal beds near the surface, and the proximity of many strip pits to populated areas. Pennsylvania anthracite occurs in four separate long and narrow fields totaling only 480 sq miles. The coal measures are rock strata and coal beds that are considerably folded and faulted. The crests of the anticlines are eroded extensively. The beds outcrop on the mountain sides and dip under the valleys. At first only the upper portions of the syn-clines could be stripped. Now stripping to increasingly greater depths is economically possible, as is indicated by the fact that the proportion of freshly mined anthracite produced by strip mining has increased from 3.7 pct of the total tonnage in 1930 to 29.6 pct in 1950. Much of the rock overlying the deeper beds now being stripped is so extensively broken that considerable difficulty is experienced in drilling satisfactory blast holes and in using explosives in such manner as to insure a uniformly broken material easily removed by the excavating machinery. Such breaking of rock strata has occurred because the bed now being stripped has been mined extensively in former years by underground methods, and tops of gangways and chambers have subsequently failed. Draglines are used to uncover coal where the overburden can be moved with little or no re-handling. These machines range in size from those having a 2 cu yd capacity bucket on a 60-ft boom to those handling a 25 cu yd bucket on a 200-ft boom. Draglines are also used to strip to the bottom of the coal basins if the depth and the distance between the crops are not too great. For this type of operation blast holes are drilled full depth to the bed. These holes are commonly 30 to 90 ft deep; however, in exceptional cases, holes may be as shallow as 12 ft or as deep as 130 ft. Drilling is normally done for blasts of 12,000 to 60,000 cu yd of overburden, 30,000 cu yd being considered an average blast if vibration is not the controlling factor. Where the stripping of wide basins or the exposure of a moderately pitching vein makes the use of draglines impractical, dipper front shovels equipped with 4 to 6 cu yd buckets load into trucks. Overburden is removed in benches of 25 to 30 ft with blast holes drilled 4 or 5 ft deeper than the planned floor of the bench. For shovels under 5 cu yd bucket capacity the volume blasted varies from 8000 to 12,000 cu yd, whereas a volume of 30,000 to 50,000 cu yd of overburden is frequently blasted at one time for the larger shovels where vibration is not an important factor. During the past decade the churn drill, generally the Model 42-T Bucyrus-Erie blast hole drill equipped for drilling 9-in. diam holes, has become the most common blast hole drilling machine. Electricity powers half the churn drills in use and is preferred on the large strippings where electric shovels are operated and the working area is concentrated. On these operations the cost of additional electricity for the drills is less than the cost of fuel to operate diesel units because of the existing large demand load of the excavating equipment. Moreover, electric motors start more easily in cold weather and generally are less expensive to maintain. Diesel driven units are employed where a higher degree of mobility is required. The average drilling speed is 8 ft per hr, although in softer rocks a rate of 15 ft per hr is attained. Where rock is hard and strata is badly broken, drill speeds may be less than 2 ft per hr. Low drilling production results under these circumstances when loose material falling from the upper portion of the drill holes causes drill stems to be jammed. Rock formations vary so greatly in the region that a 9-in. diam churn drill bit may become dull after drilling only 2 ft or may drill satisfactorily for 56 ft; however, an average of 35 ft is usual in sandstone of medium hardness. Dull bits are hoisted to flat bed trucks by the sand line of the drill and are usually sharpened in the contractor&apos;s bit shop adjacent to the job. Care is generally taken to cover the thread end of the bit with a cap. To facilitate handling of bits around the drill, a heavy thread protector having an eye top is becoming more popular than the flat-top rubber or metal cap furnished with new bits. The 9-in. diam blast holes for a 25 to 30 ft bench are normally on 18x18 ft to 20x20 ft spacings, depending on the character of the overburden, although in broken ground 15x18 ft centers may be used to obtain better breakage and a more even bottom for the bench. The patterns of holes for shots

Jan 1, 1953

By R. D. Boddorff, R. L. Ash, C. T. Butler, W. W. Kay

DRILLING and blasting in anthracite open-pit mines is a continuous problem to contractors and explosive engineers because of the diverse conditions caused by the nature of the geological formations, the extensive mining of the portions of coal beds near the surface, and the proximity of many strip pits to populated areas. Pennsylvania anthracite occurs in four separate long and narrow fields totaling only 480 sq miles. The coal measures are rock strata and coal beds that are considerably folded and faulted. The crests of the anticlines are eroded extensively. The beds outcrop on the mountain sides and dip under the valleys. At first only the upper portions of the synclines could be stripped. Now stripping to increasingly greater depths is economically possible, as is indicated by the fact that the proportion of freshly mined anthracite produced by strip mining has increased from 3.7 pct of the total tonnage in 1930 to 29.6 pct in 1950. Much of the rock overlying the deeper beds now being stripped is so extensively broken that considerable difficulty is experienced in drilling satisfactory blast holes and in using explosives in such manner as to insure a uniformly broken material easily removed by the excavating machinery. Such breaking of rock strata has occurred because the bed now being stripped has been mined extensively in former years by underground methods, and tops of gangways and chambers have subsequently failed. Draglines are used to uncover coal where the overburden can be moved with little or no rehandling. These machines range in size from those having a 2 cu yd capacity bucket on a 60-ft boom to those handling a 25 cu yd bucket on a 200-ft boom. Draglines are also used to strip to the bottom of the coal basins if the depth and the distance between the crops are not too great. For this type of operation blast holes are drilled full depth to the bed. These holes are commonly 30 to 90 ft deep; however, in exceptional cases, holes may be as shallow as 12 ft or as deep as 130 ft. Drilling is normally done for blasts of 12,000 to 60,000 cu yd of overburden, 30,000 cu yd being considered an average blast if vibration is not the controlling factor. Where the stripping of wide basins or the exposure of a moderately pitching vein makes the use of draglines impractical, dipper front shovels equipped with 4 to 6 1/2 cu yd buckets load into trucks. Overburden is removed in benches of 25 to 30 ft with blast holes drilled 4 or 5 ft deeper than the planned floor of the bench. For shovels under 5 cu yd bucket capacity the volume blasted varies from 8000 to 12,000 cu yd, whereas a volume of 30,000 to 50,000 cu yd of overburden is frequently blasted at one time for the larger shovels where vibration is not an important factor. During the past decade the churn drill, generally the Model 42-T Bucyrus-Erie blast hole drill equipped for drilling 9-in. diam holes, has become the most common blast hole drilling machine. Electricity powers half the churn drills in use and is preferred on the large strippings where electric shovels are operated and the working area is concentrated. On these operations the cost of additional electricity for the drills is less than the cost of fuel to operate diesel units because of the existing large demand load of the excavating equipment. Moreover, electric motors start more easily in cold weather and generally are less expensive to maintain. Diesel driven units are employed where a higher degree of mobility. is required. The average drilling speed is 8 ft per hr, although in softer rocks a rate of 15 ft per hr is attained. Where rock is hard and strata is badly broken, drill speeds may &apos; be less than 2 ft per hr. Low drilling production results under these circumstances when loose material falling from the upper portion of the drill holes causes drill stems to be jammed. Rock formations vary so greatly in the region that a 9-in. diam churn drill bit may become dull after drilling only 2 ft or may drill satisfactorily for 56 ft; however, an average of 35 ft is usual in sandstone of medium hardness. Dull bits are hoisted to flat bed trucks by the sand line of the drill and are usually sharpened in the contractor&apos;s bit shop adjacent to the job. Care is generally taken to cover the thread end of the bit with a cap. To facilitate handling of bits around the drill, a heavy thread protector having an eye top is becoming more popular than the flat-top rubber or metal cap furnished with new bits. The 9-in. diam blast holes for a 25 to 30 ft bench are normally on 18x18 ft to 20x20 ft spacings, depending on the character of the overburden, although in broken ground 15x18 ft centers may be used to obtain better breakage and a more even bottom for the bench. The patterns of holes for shots

Jan 1, 1952

By S. Lipson, H. W. Antes, H. Rosenthal

A study was made of the effect of ingot structure on the strength and ductility of high-strength wrought-aluminum alloys. It was found that a fine-cast structure facilitated complete homogenization which, in turn, resulted in significant increases in ductility and strength. A completely homogenized 7075-T6 alloy developed tensile properties of 85,000 psi UTS, 75,000 psi YS, with 40 pct RA. Completely homogenized 7001-T6 alloy tensile properties were 102,000 psi UTS, 99,000 psi YS, with 19 pct Ra. A method was devised for making small ingots having secondary dendrite arm spacing of less than 10 u. This method involved multiple-pass arc melting of commercial rolled plate with a tungsten electvode. This material could be completely homogenized after 3 hr at 900°F; homogenization of the original plate material was not complete after 120 hv at 900°F. Degree of homogeneity was determined by use of metallographic and electron-microprobe analyses. The electron-micro-probe study also showed the preferential segregation of solutes in the microstructure. HIGH-strength aluminum alloys, such as those of the 7000 series, usually freeze by the formation and growth of dendrites. The dendrite arm spacing (DAS) depends on the rate of solidification.&apos; Commercial ingots are usually direct chill-cast to promote more rapid solidification, but, due to the large mass of the ingot, localized solidification times are long and a large DAS results. During solidification, solute elements are rejected by the solid as it forms, causing enrichment of the liquid and ultimately solute-rich interdendritic regions. In order to attain a homogeneous ingot, the segregated solutes must diffuse across the dendrite arms. The larger the DM, the longer the time for complete homogenization. In the case of commercial ingots, the DAS is so large that the time for complete homogenization is prohibitively long and, therefore, second phases or compounds are always present. These un-dissolved phases are carried over to the wrought material during processing, resulting in an impairment of strength and ductility. In addition, the mechanical fibering of the undissolved second phases or compounds during working results in mechanical property anisotropy. If complete homogenization could be attained, higher ductility could be expected. The realization of higher ductility at current strength levels is a desirable objective; however, if higher-strength alloys were wanted, it might be possible to sacrifice some of this ductility by adding more solute elements and produce even higher-strength alloys than are currently available. Further, if complete homogenization leads to more efficient utilization of solute elements, then more dilute alloys should have relatively high strengths with very high ductility. In all instances, it would be expected that the degree of mechanical property anisotropy due to mechanical fibering would be reduced. Therefore, it was the purpose of this investigation to produce cast structures that would facilitate homogenization and to determine the effect of homogenization on the properties of high-strength, wrought-aluminum alloys. MATERIAL CLASSIFICATION Commercial Alloys. In order to illustrate the non-homogeneous condition that exists in commercial high-strength, wrought-aluminum alloys, typical micro-structures of 7001, 7075, and 7178 are shown in Fig. 1. The chemical compositional specifications of these alloys are given in Table I. It can be seen in Fig. 1 that a considerable amount of undissolved second-phase material is present in each of these alloys. The solute elements associated with the undissolved phases were identified by electron microanalyses. Back-scattered electron images and characteristic X-ray images of the three commercial alloys are shown in Figs. 2, 3, and 4. These data indicate that the second phases are regions of high copper and high iron-copper concentrations. The second-phase material also was analyzed for magnesium, zinc, manganese, chromium, and silicon, but no significant enrichment above that of the matrix was found. Therefore, the problem of homogenization resolved itself into one of dissolving the copper-rich and the iron-copper-rich second phases. In order to accomplish this objective, two approaches were made. The first was to reduce the iron as low as possible since this element has a maximum solid solubility of 0.03 pct in aluminum. The second was to produce cast structures with finer DAS to facilitate dissolving the second phases. Commercially Produced High-Purity Alloys. A special high-purity, 2000-lb ingot of 7075 alloy was made by a commercial producer. This alloy contained the following weight percentages of solutes: 5.63 Zn, 2.48 Mg, 1.49 Cu, and 0.21 Cr. All other elements combined were less than 0.02 pct by wt including iron and silicon at less than 0.01 pct each. The ingot was cast and processed into rolled plate using standard commercial techniques. Microstructures of standard commercial 7075 and the special high-purity 7075 are shown in Fig. 5. It can be seen from this figure that the high-purity alloy has less undissolved second-phase material, but a significant amount was still present. The second phase in the high-purity material did not contain iron but it was found to be enriched with copper. The slight effects of the increased purity and de-

Jan 1, 1968

By G. Ansel, J. P. Doan

THE important literature concerning zirconium in magnesium-base alloys is predominantly contained in patent references. Sauerwald, Eisenreich, and Holub1-4 discovered the profound grain-refining influence of small amounts of zirconium on cast magnesium and the attendant increases in strength, and ductility. They extended their work to reveal that limited amounts of certain other elements, including zinc, could be added for further improvement in strength and hardness. British. Patent 5111376 relates to the parent discoveries of Sauerwald, Eisen¬reich, and Holub. All of these references deal primarily with casting characteristics and properties in the cast state. Minor mention is made of desirable tensile strength, ductility, and toughness of wrought Mg-Zr alloys* modified by one or more additional elements. No extensive reference in the literature" has been found to the particular characteristics of wrought Mg-Zn-Zr alloys, with which the present paper is concerned. - The work to be discussed represents essentially an extrusion study of the Mg-Zn-Zr alloys and indicates that this system is highly interesting in the wrought state both academically and commercially. Accordingly, an important aim of this paper is to describe the hot-workability, tension and compression properties, metal¬lography, and toughness of these alloys. Compositional variations are limited to the ranges o to 7 per cent Zn and o to I per cent Zr. EXPERIMENTAL PROCEDURE Alloying and Casting Alloying ingredients were cell mag¬nesium, commercial zinc (intermediate grade), and a chloride salt containing approximately 50 per cent ZrCl2 and 50 per cent alkali chlorides. This salt is produced by Titanium Alloy Manu¬facturing Company. Melting, alloying, and pouring procedure essentially followed the crucible process as described by Nelson.6 The magnesium was melted using Dow No. 310 flux, and the temperature raised to 1350° to 1400°F for the alloying operations. After alloying of the zinc, the zirconium containing salt was added in small portions with vigorous stirring. This is an adaptation of the method described in British Patent 533264.7 The alloying efficiency of zir¬conium was approximately 30 per cent. During the alloying of zirconium, fluorspar was added in amount required as a flux¬inspissating agent. Next the melt was refined with Dow No. 310 flux and held quietly for 15 min. Finally billets were cast in a book-type permanent mold

Jan 1, 1946

By O. W. Simmons, L. W. Eastwood, C. M. Craighead

The results of a preliminary study of 113 ternary titanium-base alloys are described. The compositions investigated were as follows: 1. Ternary titanium-carbon alloys containing copper, silicon, vanadium, chromium, manganese, iron, or cobalt. 2. Ternary titanium-nitrogen alloys containing chromium. 3. Ternary titanium-chromium alloys containing additions of vanadium, molybdenum, tungsten, cobalt, or nickel. 4. Ternary titanium-manganese alloys containing additions of silicon, chromium, tungsten, or iron. Tensile properties, minimum bend radii, hardnesses, response to heat treatment and aging treatment, and phase relationships for these alloys were determined. THIS is the second in a series of papers describing some of the results of the exploratory work on titanium alloys being conducted at Battelle Memorial Institute for the Wright-Patterson Air Force Base. The work with the binary alloy systems has been described in a previous publication.&apos; This paper describes ternary titanium-base alloys as follows: 1. Ternary titanium-carbon alloys containing copper, silicon, vanadium, chromium, manganese, iron, or cobalt. 2. Ternary titanium-nitrogen alloys containing chromium. 3. Ternary titanium-chromium alloys containing additions of vanadium, molybdenum, tungsten, cobalt, or nickel. 4. Ternary titanium-manganese alloys containing additions of silicon, chromium, tungsten, or iron. As the preparation and fabrication of the 0.5-Ib experimental titanium alloy ingots to 14-ga sheet, as well as the methods of testing used in this work, have been described in the first paper of this series, Titanium Binary Alloys, no further discussion of these techniques will be given. Ternary Alloys Containing Carbon The following alloys of titanium, containing carbon and one other metal, were investigated: 1. 0.25 pct carbon; 1, 2.5, and 5 pct copper. 2. 0.25 pct carbon; 1, 2, and 3 pct silicon. 3. 0.25 pct carbon; 1, 2.5, and 5 pct vanadium. 4. 0.25 pct carbon; 2.5, 3.5, 5, 10, and 15 pct chromium. 5. 0.5 pct carbon; 2.5, 3.5, and 5 pct chromium. 6. 0.25 pct carbon; 1, 1.75, 2.5, 3.5, 5, 7.5, and 10 pct manganese. 7. 0.5 pct carbon; 1.75, 2.5, 3.5, 5, 7.5, and 10 pct manganese. 8. 0.25 pct carbon; 0.25, 0.5, 1, and 2 pct iron. 9. 0.25 pct carbon; 1, 2, and 3 pct cobalt.

Jan 1, 1951

By B. Chalmers, Stefan Wojciechowski

Rotary oscillation of the mold has been used to produce mechanical stirring of the melt during unidirectional solidification. Al-Cu alloys (1 to 4 wt pct Cu) were solidified to study the influence of alloy composition, temperature at which intensive cooling was initiated, and intensity of stirring on the temperature distribution in the sample, the resulting structure, and the copper microsegregation at the columnar to equiaxed transition. The Proposed explanation for the equiaxed solidification in stirred castings consists of the melting off of dendrite fragments, their partial remelting in hotter liquid, and the subsequent growth of the surviving part when the temperature of the melt is lowered. This may result in the formation of a crystal network, which can stop the growth of the columnar crystals and produce the equiaxed zone. It has long been known and many times affirmed that ultrasonic vibrations or relative motion of liquid and solid can significantly reduce the length of the columnar zone and refine the grains in the equiaxed zone in ingots and castings. Because of the technological importance of these phenomena many investigations have been conducted in which ultrasonic energy or mechanical vibration,1-6 electromagnetic stirring,7"10&apos;17 magnetic stirring,"&apos;12 or mechanical stirring13&apos;14 were used to produce the relative liquid-solid motion. More recently, the question of the importance of natural (or thermal) convection has been raised.15-l7 Several possible mechanisms have been suggested.l,8,18-20 They can be divided into two main groups. The first group includes such nucleation hypotheses as nucleation frequency enhancement due to cavitation at the mold wall or other surfaces; cavita-tion in the body of the liquid resulting from the pressure wave, or an agglomeration mechanism, which involves the formation of nuclei by the accretion of subcritical nuclei. The second group considers the grain-refining action to be due to crystal multiplication phenomena such as the removal of the dendrite fragments by viscous forces or by a high-pressure cycle following cavitation or by a melting-off process. Sbme recent investigations give good experimental support to the melting-off proposa1.21-23 The purpose of the present study was to examine the effect of mechanical stirring during the unidirectional solidification on the structure of alloy samples and on the solidification variables. The motion of the liquid was produced by rotary oscillation of the mold around its vertical axis.

Jan 1, 1969

By John H. Hollomon

THE interpretation of notched-bar impact results has been a matter of controversy since the introduction of more or less standard tests by Fremont,1 Charpy2 and others at the turn of the century. Many investigators3-l1 have contributed to the understanding of the significance of such tests. Several symposiums12,13 have been held with the express purpose of discussing the significance and interpretation of the impact tests. It is not the purpose of this paper to review all these contributions to the development of the knowledge on this subject, but rather to present an interpretation of notched-bar impact tests that appears to be in agreement with the available published data. Fundamentally, the interpretation discussed in this DaDer of the brittle failure of some steels in notched-bar impact test is similar to the qualitative analysis presented by Ludwik3 in the 1920&apos;s. Essentially, Ludwik stated that the function of the notch was to increase the tensile stress necessary for yielding, and that brittle failure occurred when the stress necessary for yielding exceeded the normal stress required for fracture (with no plastic de- formation). Ludwik, however, had no means of measuring this normal stress necessary for failure of a plastically* undeformed metal, therefore he could not extend his interpretation beyond the qualitative stage. Kuntze14 and recently MC- Adam15 attempted to determine the effect of deformation on the stress necessary for fracture by testing, in tension, specimens notched to various depths and angles. Because of the uncertainties in Kuntze&apos;s (and RlcAdam&apos;s) analysis of the stresses (which have been pointed out by SachsI6) and the inherent nonuniformity in the stress distribution in notched bars, the results of such investigations are difficult to interpret. A more direct method of measuring the effect of deformation on the fracture stress was suggested in previous papers17-l8 and has been utilized in the present investigation. The present paper concerns itself primarily with the behavior of notched impact specimens. The effects of strain rate, temperature, and stress distribution, which are discussed with reference to the impact specimen, apply just as well to the behavior of metal at the bases of notches in any engineering structure. All that is necessary is to determine the stress distribution in the structure at the base of the notches.

Jan 1, 1944

By Ernest Klepetko

A notable amount of fossil resin exists in many of the bituminous coal beds of Utah, The upper part of these show a marked concentration of resin, which occurs primarily in the fracture seams. In general, these seams are less than 1/8 in. thick, but it is not uncommon to find coal with areas as thick as 3/8 in. The resin is of the hard kauri variety, with coloration varying from deep blood red to lemon yellow and almost colorless. There is also an almost black variety, which is not readily soluble in the same solvents as the lighter colored resins. These resins are valuable for many industrial purposes. When refined, they are suitable for uses similar to those for which lac resins are usually employed. They are particularly desirable for preparing dielectric varnishes. They are also desirable in preparing printing inks, waterproofing agents, special synthetic rubbers and rubber cements, They are especially suitable as an admixture in synthetic rubbers and cements, as they supply the essential tackiness the synthetics lack. The following chemical trade characteristics&apos; have been found by Inter-Chemical Co. research chemists in the refined resins: Specific gravity (melted)....... I. 03 -1.06 Acid number.................. 6-8 Softening point, mercury method 160°-165°C Iodine valus (Wijs)............ 140-150 Refractive index............... 1.544 Color (Hellige, 25 pct toluol)... 14 Molecular weight, average...... 732 purity, 99 + pct pure hydrocarbon "Unaffected by alkalies and resistant to one hour fusion with solid potassium hydroxide. Completely soluble in aliphatic and aromatic hydrocarbons. Insoluble in alcohols, ethyl acetate and other commonly used alcohols and esters. Compatible with natural and synthetic rubbers, chlorinated and cyclized rubbers, waxes, vegetable oils, mineral oils, rosin, ester gum, phenolic and modified phenolic resins and other natural and synthetic resins of like solvent solubility."* The recovery and refining of these fossil resins has been the subject of much research work and has intrigued many investigators. Broadly speaking, the problem involves separation of the resin from the coal into a very high-grade concentrate, and subsequent treatment of this concentrate to produce an acceptable refined resin for its various usages. Obviously it is also desirable to recover the residua] coal without excessive size degradation. The late W. D. Green, concentration engineer for Combined Metals Reduction CO., obtained U. S. Patent No. 1773997 for the recovery of resin from coal through

Jan 1, 1949

By Richard E. Goodman

Rock slope stability assumes different roles in decision making as a mining venture develops and, accordingly, geological investigations vary in thought and in deed according to the project stage. During exploration, stability questions bear on the decision whether or not to mine, because their resolution affects the cost of mining. During the mine lay-out stage, stability is a key factor in deciding how to mine. Finally, during mine operation, stability affects safety and operating costs, and therefore, pit productivity. Among the many considerations influencing stability, rock mass classification and geological environmental factors count heavily. The rock mass behavioral class depends upon the rock character and its degree of fracturation. The loads on the rock are derived from the environmental factors of initial state of stress and the water regime. Of course, stability is also tied to a host of non-geological factors such as the pit geometry, mining style, relative wealth, philosophy of mine operation, and accident history. METHODS OF STABILITY EVALUATION AND CORRESPONDING INVESTIGATORY TOOLS The purposes of the geological investigation are to identify and evaluate the geological factors influencing stability so that the vital questions can be analyzed. The tools selected for stability analysis reflect the stage of development of the mining scheme and therefore the geological investigations, which must furnish data for the analyses and must vary with the project stage. For exploration and mine planning, the methods of resolving stability questions must meet the need to evaluate numerous parameters in limited time; examples of such methods are stereographic projections (1, 2, 3) charted solutions(4) and simple physical models (5, 6). An investigation program that employs geological mapping tools and aids to mapping, such as geophysics, with relatively few rock tests, is in balance with these analytical methods. First data about the water regime are obtained from observed piezometric levels, and the initial state of stress is queried simply through geological inference and experience. In the layout stage more elaborate physical scale models (7) and perhaps finite element models (8, 9, 10, 11, 12), will be relevant to the design process and therefore much more detailed geological investigation is required. The scale of observation of both rock material and joints will be significantly enlarged and mechanical tests of rock and joint strength

Jan 1, 1972

By Ernest Klepetko

A notable amount of fossil resin exists in many of the bituminous coal beds of Utah. The upper part of these show a marked concentration of resin, which occurs primarily in the fracture seams. In general, these seams are less than 1/8 in. thick, but it is not uncommon to find coal with areas as thick as 3/8 inches. The resin is of the hard kauri variety, with coloration varying from deep blood red to lemon yellow and almost colorless. There is also an almost black variety, which is not readily soluble in the same solvents as the lighter colored resins. These resins are valuable for many industrial purposes. When refined, they are suitable for uses similar to those for which lac resins are usually employed. They are particularly desirable for preparing dielectric varnishes. They are also desirable in preparing printing inks, waterproofing agents, special synthetic rubbers and rubber cements. They are especially suitable as an admixture in synthetic rubbers and cements, as they supply the essential tackiness the synthetics lack. The following chemical trade characteristics have been found by Inter-Chemical Co. research chemists in the refined resins: Specific gravity (melted)....... I .03 -I.06 Acid number.................. 6-8 Softening point, mercury method 160º-165ºC Iodine value (Wijs)............ 140-150 Refractive index............... 1.544 Color (Hellige, 25 pct toluol)... 14 Molecular weight, average...... 732 Purity, 99 + pct pure hydrocarbon "Unaffected by alkalies and resistant to one hour fusion with solid potassium hydroxide. Completely soluble in aliphatic and aromatic hydrocarbons. Insoluble in alcohols, ethyl acetate and other commonly used alcohols and esters. Compatible with natural and synthetic rubbers, chlorinated and cyclized rubbers, waxes, vegetable oils, mineral oils, rosin, ester gum, phenolic and modified phenolic resins and other natural and synthetic resins of like solvent solubility."* The recovery and refining of these fossil resins has been the subject of much research work and has intrigued many investigators. Broadly speaking, the problem involves separation of the resin from the coal into a very high-grade concentrate, and subsequent treatment of this concentrate to produce an acceptable refined resin for its various usages. Obviously it is also desirable to recover the residual coal without excessive size degradation. The late W. D. Green, concentration ennineeru for Combined Metals Reduction CO., obtained U, S. Patent No. 1773997 for the recovery of resin from coal through

Jan 1, 1948

By Ernest Klepetko

A notable amount of fossil resin exists in many of the bituminous coal beds of Utah. The upper part of these show a marked concentration of resin, which occurs primarily in the fracture seams. In general, these seams are less than 1/8 in. thick, but it is not uncommon to find coal with areas as thick as 3/8 inches. The resin is of the hard kauri variety, with coloration varying from deep blood red to lemon yellow and almost colorless. There is also an almost black variety, which is not readily soluble in the same solvents as the lighter colored resins. These resins are valuable for many industrial purposes. When refined, they are suitable for uses similar to those for which lac resins are usually employed. They are particularly desirable for preparing dielectric varnishes. They are also desirable in preparing printing inks, waterproofing agents, special synthetic rubbers and rubber cements. They are especially suitable as an admixture in synthetic rubbers and cements, as they supply the essential tackiness the synthetics lack. The following chemical trade characteristics have been found by Inter-Chemical Co. research chemists in the refined resins: Specific gravity (melted)....... I .03 -I.06 Acid number.................. 6-8 Softening point, mercury method 160º-165ºC Iodine value (Wijs)............ 140-150 Refractive index............... 1.544 Color (Hellige, 25 pct toluol)... 14 Molecular weight, average...... 732 Purity, 99 + pct pure hydrocarbon "Unaffected by alkalies and resistant to one hour fusion with solid potassium hydroxide. Completely soluble in aliphatic and aromatic hydrocarbons. Insoluble in alcohols, ethyl acetate and other commonly used alcohols and esters. Compatible with natural and synthetic rubbers, chlorinated and cyclized rubbers, waxes, vegetable oils, mineral oils, rosin, ester gum, phenolic and modified phenolic resins and other natural and synthetic resins of like solvent solubility."* The recovery and refining of these fossil resins has been the subject of much research work and has intrigued many investigators. Broadly speaking, the problem involves separation of the resin from the coal into a very high-grade concentrate, and subsequent treatment of this concentrate to produce an acceptable refined resin for its various usages. Obviously it is also desirable to recover the residual coal without excessive size degradation. The late W. D. Green, concentration ennineeru for Combined Metals Reduction CO., obtained U, S. Patent No. 1773997 for the recovery of resin from coal through

Jan 1, 1948

By L. G. Huntley

The entire production of Colombia for 1927 was that from the Tropical Oil Co. concession. Production was approximately 14,000,000 bbl., of which 12,081,000 bbl. were exported from the Mamonal terminal of the Andian National Corp. near Cartagena. The balance represents that refined and used or marketed within Colombia. As production in 1926 was only 6,443,540 bbl., the increase has been more than 100 per cent. This is partly due to the looping of the pipe line from the concession to the coast, and the resultant increase in capacity to 50,000 bbl. per day, and also to increased domestic sales. The refinery at Barranca has been doubled in capacity, and it is understood is now handling 5000 bbl. of crude daily. This gives the producing field a potential outlet of 55,000 bbl. daily. The Tropical Oil Co. completed 82 wells in 1927, with one abandoned for drilling difficulties. The average daily initial production was 1136 bbl. per well. The largest well is reported to be No. 173 La Cira, rated as 5000 bbl. initial. Most of the wells were drilled on the Infantas and La Cira structures, the latter located en echelon to the northwest of the main Infantas anticline, and separated from it only by a low saddle. Drilling operations on the Llanos fold, also en echelon with Infantas but to the southeast, and at San Luis, bearing about the same relationship to the Umir Hills uplift, are still more or less in the exploratory stage. The four structures mentioned cover approximately 15,000 acres, on which there are at present 223 producing wells. Drilling has been better than 92 per cent. successful since starting operations on this concession. Allowing 15 acres per well, which is 50 per cent. more than good practice in the Mid-Continent fields, there is thus room for 1000 wclls, which at the recent rate is a 10-year drilling campaign. This does not take into consideration the known deeper oil zones, which have not been prospected, but which are uniformly petroliferous at their outcrops to the eastward. It is estimated that with the enlarged pipe line capacity, as well as increased domestic consumption, the production for this country will be about 20,000,000 bbl. for the current year of 1928. It is possible that

Jan 1, 1928

Discussion of the paper of E. C. HARDER, presented at the Pittsburgh meeting, October, 1914, and printed in Bulletin No. 94, October, 1914, pp. 2573 to 2586. I. C. WHITE, Morgantown, W. Va.-I have seen some of the great iron-ore deposits described in this paper. -One of these in the State of Minas Geraes appears to be an immense mountain of iron ore about 2,000 ft. in height. All over its surface from bottom to top may be seen large masses of rich iron ore, varying in weight from a few pounds to several tons, and from all that one can judge without exploitation, the quantity of available ore is certainly very large in this particular mountain. Whether it will prove a mere surface deposit or shell, like the famous Iron Mountain of Missouri, is for future exploration to determine. The question of transportation is the main factor in determining the availability for exportation of these Brazilian iron ores. There is no coal in Brazil that can be manufactured into coke without expensive previous treatment, since the best of the raw coal in the States of Rio Grande do Sul, Santa Catharina, and Parana, contains about 35 per cent. of ash, of which about 6 per cent. is sulphur, and coal of this composition cannot be utilized in iron smelting without purification. When I was chief of the Brazilian Coal Commission in 1904 to 1906, I had a cargo of this coal shipped to Balk, Germany, and treated in the great coal-testing plant of the Humboldt Engineering Works, located just across the Rhine from Cologne. It was found that by crushing and washing the Brazilian coal, it yielded 33 per cent. of briquetting coal, containing only 10 to 12 per cent. of ash (of which 1 per cent. was sulphur), and 42 per cent. of slack coal, with 25 per cent. of ash and low in sulphur, so that about 33 per cent. of the Brazilian coal could be converted into coke of fair quality, as shown by the following average of three analyses, using air-dried coal, of briquets made from the Barro Branco coal bed, State of Santa Catharina: Moisture 1.56 Volatile matter 31.66 Fixed carbon 56.73 Ash : 10.05 Total 100.00 Sulphur 1.34 Phosphorus 0.003 B.t.u 13,395

Jan 4, 1915

By James A. Beck

Phosphorus is essential to all life processes and therefore to the existence of man. In this role, there are no substitutes for phosphorus. In a commercial sense, phosphorus and its compounds are important to our entire economy, both in agricultural and in nonagricultural applications. The world&apos;s phosphate requirements are met by mining phosphate rock. Two general types of phosphate ore deposits are commercially important-igneous and sedimentary. Igneous deposits generally are more difficult and more expensive to mine than sedimentary deposits and are of importance mainly in the USSR, North Vietnam, and South Africa. Approximately 15% of current world production comes from these deposits. Over 80% of world production is from the pellet sedimentary deposits of Florida, North Africa, and elsewhere. Guano, the second type of sedimentary deposit, is produced in South America but is declining in world importance. Production Production of phosphate rock was reported in 31 countries during 1972. However, about 80% of world output is produced in three countries: the United States, the USSR, and Morocco (Table 15.2A.1). Throughout the past 20 years, between 75% and 80% of world output has been produced by these three leading countries. Between 1948 and 1972 world production increased at an average rate of 10% per year, from 21.5 to 103.9 million tons. Growth was quite steady at 6% per year through 1963, but in the 5 years from 1963 to 1968 production increased 11 1/2% per year. Since 1968 production has increased about 91/2 % annually. Since 1948 several new phosphate mining ventures have been developed in Senegal, Togo, Israel, and North Vietnam. In the same period, only one important phosphate deposit, on Makatea Island, has been mined out and closed. A number of other phosphate mining ventures are in various stages of planning and development. The most important new venture in phosphate mining is in the Spanish Sahara. New mining capacity is expected also in Syria, and serious investigation is in progress in Peru, India, Turkey, and Australia. Major expansion of existing operations is expected in Morocco, Russia, Tunisia, Israel, and Jordan. Consumption Historical data on world consumption of phosphate rock are not available in detail, but world consumption of phosphate has closely paralleled world production in the past two decades. The United States is the world&apos;s largest user of phosphate rock, accounting for 30%

Jan 1, 1976

By B. H. Bergstrom, Will Mitchell, T. G. Kirkland, C. L. Sollenberger

IN a previous article&apos; the authors outlined a study of the variables in rod milling and also reported data from a series of open circuit grinding tests on a massive limestone in a 30-in. x 4-ft end peripheral discharge rod mill. As a second part of the experimental program, an analysis is now presented for the 30-in. x 4-ft overflow rod mill grinding under identical conditions, except that discharge ports on the periphery of the mill shell have been sealed so that the products from the present series overflowed through a 9-in. diam .opening in the center of the end plate. A variance analysis has been made of the combined data for the two experiments, and performances of the two mills are compared here. Included in the first report&apos; were descriptions of feed preparation, rod mill circuit, instrumentation and controls, and techniques used to evaluate data. Dependent and independent variables were defined, and variance analyses were made to test the relative significance of variables and to establish magnitude of error for the experiment. Significant data were plotted in various combinations, and conclusions were drawn from the graphs. The procedure and analysis in this series of tests follows the first tests and is not repeated. Data from the second series are recorded in Table I. Listed in the first three columns are the independent variables of feed rate (1000, 2000, 3000, 4000, and 5000 1b per hr), mill speed (50, 60, 70, 80, and 90 pct of critical), and pulp density (50, 60, 70, and 80 pct). The dependent variables, Pso, P100, reduction ratio, slope of the log-log sieve analysis curve, power demand, and Bond work index follow. Of these, only the reduction ratio and the Bond work index were analyzed for significance. Production of new surface as calculated from sieve analyses has not been included for this series because of the questionable assumptions that have to be made to satisfy the formulas involved. The large number of products obtained during the runs precluded the use of surface measurement techniques by the gas adsorption methods at this time; however, samples of all products have been stored for future reference. To test the consistency of the reporting of the sieved products, an averaged sieve analysis was calculated from the wet-dry plots obtained from the three product samples of each run. The resulting averaged analysis was plotted and the P80, selected. The relative deviations of the P80&apos;s from each of the three product samples with respect to the P80 of the averaged analysis were then calculated. In only two sets were the relative deviations (6.2 and 9.9 pct) considered excessive. In each of these two sets, one sieve analysis was obviously out of line; hence that analysis was ignored and new averages were computed. This reduced the relative deviations to 1.2 and 2.7 pct respectively. The relative deviations of the product analyses with respect to their averages ranged from 0.1 to 1.4 pct at 1000 lb per hr, 0.0 to 1.1 pct at 2000 lb per hr, 0.2 to 3.0 pct at 3000 lb per hr, 0.3 to 4.3 pct at 4000 lb per hr, and 0.5 to 5.2 pct at 5000 lb per hr. The relative deviation of the 80 pct passing point for 96 dry sieve analyses of the feed with respect to that of the averaged analysis was 7.6 pct. This slightly higher percentage can probably be attributed to a greater proportion of tramp oversize in a crusher product than is ordinarily found in a rod mill product. The last column on Table I lists the adjusted work index, which has been used as the measure of efficiency for the various combinations of operating conditions investigated. Efficiency increases as the index becomes lower. It was reported in the previous paper that the work indexes for the Waukesha limestone used in these experiments decreased as the product size decreased (as calculated from Bond grindabilities). That is, this limestone becomes easier to grind as the material becomes finer. This is unusual, because the work index for most materials as calculated from the Bond grindability has remained constant as the product size decreased or has increased slightly. Table II lists the results of Bond grindability tests at all mesh sizes from 3 to 200 and the work indexes calculated from them. To remove this variation of work index with product size from the data so that results would apply to any material of constant work index, the work index values shown in Table II were plotted against product size on log-log paper. From this curve (a straight line function in this case), the expected work index for the product size for each of the runs of the experiment was obtained. The work indexes as calculated from the reduction ratio and energy consumption were then divided by the corresponding expected work index. The results obtained are reported in percentages on Table I as adjusted work index and are actually percentages of the work index for the Waukesha limestone at the size in question. Multiplication of the work index value for a material of constant index by these percentages should allow the application of the adjusted work index curves to the material. Only the adjusted work index values, not the actual experimental values, were used for the variance analyses and for the graphs.

Jan 1, 1956