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By G. F. Raymond

Discussion by G.F. Raymond Knudsen's study presents two curious conclusions: • The kriging of blasthole assays can systematically overstate mill head grades by as much as 21% as a result of unbiased sample variance. • No estimation method is able to reduce this bias to a small margin. The study is based on a simulation using real data and what are presumed to be actual variogram parameters from a real deposit. Although I have no doubt that the first conclusion (21% overestimation) is correct for the author's simulation, I do not believe this represents a realistic mining situation. Over the past 15 years I have done extensive comparisons between exploration drill-hole assays, blasthole assays and mill head grades on seven major open-pit mines, including some very erratic gold deposits. Commonly, nugget effects on blasthole variograms were 10% - 20% higher than on exploration variograms. And in one extreme case, the difference was 50%. Even in the extreme case, ordinary kriging on blastholes agreed well with the mill head grade over the long term. In Knudsen's simulation, the blasthole nugget effect is assumed to be 200% higher than exploration data's. He supports this by variogram plots from each. My guess is that the apparent, large difference between these variograms results from a failure to account for the proportional effect (blasthole assays are likely from a higher-grade area). A simple check would be a comparison of the variance of exploration samples nearest blastholes. As for a nearly conditionally-unbiased estimator of a large random error, the arithmetic mean of all of the data certainly qualifies, provided there is an even data spacing. As a corollary, so does simple kriging, which would include, in this case, a large weighting to the arithmetic mean. Similarly, using a large number of samples with ordinary kriging or indicator kriging would significantly reduce the bias in the case of a large nugget effect for the variogram.[ ] Reply by H. Peter Knudsen Raymond questions two conclusions in the paper. First, he wonders whether a 21 % overestimation represents a realistic mining situation. I agree that 21 % is high, but overestimation in the range of 10% to 15% is certainly common in my experience. Furthermore, several years ago I consulted on a gold mine that was experiencing a 45% overestimation due, predominantly, to poor blasthole samples: In further questioning of the 21 % value, Raymond wonders whether the nugget effect of the blastholes is really so much larger than the nugget values of the exploration data. It is consistently larger throughout the deposit. In my experience with six Nevada gold mines, the high nugget value is not unusual. In fact, for some reason, nugget values for blasthole samples are typically about 0.0005 (opt squared). I am of the opinion that this high nugget effect observed at many gold mines is predominantly due to the inherent inadequacies of the blasthole sample and subsequent sample preparation. The second conclusion Raymond questions is the inability of the estimators tested to reduce the conditional bias. In fact, the conditional bias is extreme with the polygon estimator and greatly reduced by ordinary kriging. However, it was not eliminated. Raymond suggests using simple kriging, or perhaps a larger number of samples, to reduce the conditional bias. The technique of simple kriging may be less affected by the random errors in the data, but I did not test the technique. Using a larger amount of data presupposes that too few samples were used initially. In ordinary kriging and indicator kriging, the screen effect comes into play and ensures that samples beyond the second screen are given zero weight. Hence, increasing the sample size does not change the estimates nor the conditional bias. The main point of my paper is that the random unbiased errors (a fact of life in blasthole samples) cause a conditional bias in our estimates. The mechanics by which the conditional bias is introduced are nicely explained by Springett. My paper simply shows that the bias is also present when working with linear estimators, such as ordinary kriging, and even with nonlinear estimators, such as indicator kriging.[ ]

Jan 1, 1993

By R. Dawes, B. R. Davis, T. LeBel, C. Pelow

Settling of copper, lead, and other metals in slags is an important part of overall recovery of metals, both in primary and secondary operations. The impact of slag reduction on settling is not well known. This is important since the cycle time for slag cleaning can be significantly changed with reduction of slag, through a decrease in slag viscosity and growth of metal/sulphide particles. KPM has worked with a number of primary and secondary recyclers to better understand how reduction of slag can improve overall operation. In this presentation, we describe the type of test work and analysis done and how this information can be used to guide improvements in metal recovery from slags. INTRODUCTION Typically, copper smelter slags contain about 1-2% copper, both in the form of copper and copper sulphide beads and dissolved copper oxide in the slag (converter slags are recycled to the smelter). With the average grade of world copper deposits close to 0.7%, it is practical to recovery copper from these slags. This is done routinely by one of two methods. The first method uses slow cooling in large ladles or beds to grow particles of copper sulphide. The slag is then crushed, milled, and floated for copper recovery. The slow cooling promotes precipitation and growing of the sulphide particles from the solution and then, the slag can then be milled to recover the copper through froth flotation. A consideration of this process includes the energy cost related to grinding, the use of water, capital cost for ladles, and the availability of tailing dumps. At the same time, the retreatment of the Cu concentrate produced in the mill will displace fresh concentrate, which ultimately results in a reduction of the throughput. The second method is the so-called “slag cleaning” and is based on the settling of copper and copper sulphides particles dispersed in the slag. This is achieved by decreasing its viscosity by increasing the slag temperature and reducing the iron spinel solid phases (magnetite) that increase the apparent viscosity of the slag. Typically, the reduction does not extend to the reduction of copper, and soluble copper (mainly as oxide) and fine matte particles remain at concentrations typically at or above 0.5% (Cardona et al., 2011; Cardona, 2011). This is an improvement on 20 years ago when slag was sent to dumps at 1% or less (Demetrio et al., 2000).

Jan 1, 2019

By W. F. Brace, James D. Byerlee

Although prediction of failure of rocks may be ultimately possible without knowing in detail how failure occurs, understanding of this process is certainly of fundamental interest in rock mechanics. Of the various failure mechanisms, probably brittle fracture is of greatest interest. Recent reviews have brought this subject nearly up to date (Hock,1 Price,2 and Jaeger 3) and here, we summarize certain recent developments. As our understanding of brittle fracture is derived almost solely from laboratory studies this will amount to a review of certain laboratory experiments. New studies in four areas are pertinent; these concern 1) the volume changes which accompany brittle fracture, 2) the propagation of cracks in compression, 3) certain frictional characteristics of rocks and minerals at high pressure, and 4) the law of effective stress for crystalline rocks. Our objective here is less to present an entirely consistent, complete picture of how a rock fractures than it is to emphasize gaps in our understanding and, thus, suggest areas that need further study. Progress is still painfully slow in this field and perhaps this approach will stimulate new work. OBSERVATIONS Although the behavior of large rock masses is most often the concern in rock mechanics, the number of large-scale experimental tests is still vanishingly small. Our experimental observations arc, by and large, still made on small, relatively flaw-free, rock samples. Here also we report the behavior of small samples.

Jan 1, 1967

By R. W. Heckel

The conzpaction of metal powders is analyzed through density-pressure curves as a three-stage process — die filling, individual particle motion, gross compact deformatim. The densification occurring during each of these stages is evaluated quantitatively. The amount of densification in the first and second stages is primarily a function of particle geometry. The densification taking place in the third stage is dependent mainly upon the material. AT the present time the densification of powders during compaction is usually described by terms such as "compactibility," "compressibility," and "compression ratio." Schwarzkopfl has proposedthat compactibility be defined as the minimum pressure needed to produce a given green strength, while compressibility be used to indicate the extent towhichthe density of a powder is increased by a given pressure. Compression ratio is generally defined as the ratio of the compact density obtained at a given pressure to the apparent density of the loose powder. 13alshina found that the pressure imposed on a compact affects the relative volume of the compact according to the following relation: hP= - LV + C [ll where: P is the applied pressure, V is the relative volume of the compact, Ratio of the specific volume of the compact to that of the material without porosity. Lisa constant described as the modulus of pressing and is determined by the slope of a plot of h P vs V, and C is a constant which is determined by the intercept of a plot of InP vs V. Several drawbacks of the Balshin analysis may be noted from Fig. 1, which is a plot of pressure (loga-

Jan 1, 1962

By Peter N. Calder, Garston Blackwell

"The history and analysis of a structurally complex open-pit rock slope displacement are described. The failure involved part of the only haulage access to a major ore zone. The ongoing movement was monitored using an electronic distance measuring device and a continuous electronic system specifically designed for the problem. A major fault, infilled with a thick clay gouge, and intersected by other structural features, resulted in the transfer of weight to a rock mass acting as a buttress in the lower regions. Movements were affected by blasting vibrations and ground-water conditions. Remedial measures included limiting blasting levels and lowering the water table, following which mining was successfully resumed in the area.IntroductionBrenda Mines is situated in a mountainous region of the southern interior of British Columbia; the location is indicated on Figure 1 . The elevation is approximately 5000 ft , giving a climate typical of more northern areas. Temperatures vary from -30°C to + 30°C. A snowfall level of 400-800 cm is experienced from November to April. During the spring, a heavy runoff period occurs, increasing the pumping rate in the pit from a nominal 50* gal/min to 2000 gal/min.The orebody occurs in a fractured Jurassic rock mass referred to as the Brenda Stock; it is approximately 3000 ft in diameter and 1000 ft deep. The Brenda Stock was intruded into the Upper Triassic Nicola Group tuffs, breccias and argillites. Mineralization occurred in several stages in a relatively homogeneous quartz diorite host rock. Shearing, faulting and jointing occured later in many stages, producing mineralized clay gouge zones up to several feet in thickness'> , These gouge zones have a major effect on the stability of the pit walls.The structural geology of the pit walls is otherwise relatively simple. Figure 2 is a conventional pole diagram indicating the orientation of the rock structural systems which are present. The major system has an average dip of approximately 70 degrees to the south and strikes approximately east /west , it is not undercu t by the pit slope. Figure 3 is a photograph of part of the North wall on which this system is indicated.The basic pit design consists of a working bench height of 50 ft, with a final wall height of 100 ft between berm s and a 100-ft-wide berm every second lift. Combined with a local face slope of 50-70 degrees, the resulting wall slope, not including haulage roads, is 45 degrees. Figures 4 and 5 are generalized plans and sectional views of the pit."

Jan 1, 1980

By Murray C. G

A database of over 1300 earthquakes has been used to investigate the relationship between seismicity and tectonics in two areas of eastern Queensland and NE New South Wales. These areas have been monitored by several networks that were installed primarily to monitor seismicity associated with the impounding of large reservoirs. Events that occurred prior to the installation of these networks have been excluded from the analysis because they have large uncertainties in location.Seismicity in SE Queensland and NE New South Wales is concentrated in the coastal regions. Activity also extends off the coast of Queensland and beyond the continental shelf. The coastal activity is mainly within the New England Fold Belt. An area of the fold belt overlain by the Clarence-Moreton Basin is also active. There is an area of activity near St. George that correlates to the Wunger Ridge. Fraser Island and the adjacent onshore area, near Maryborough, are notably aseismic.Within the New England Fold Belt there is a NNW trending zone of concentrated activity that includes the South D'Aguilar Block and the Esk Trough as far as Gayndah. The Nambour Basin, the Gympie Province, and the southern portion of the Yarrol Block are significantly less active.Composite focal mechanisms of several groups of accurately located earthquakes in SE Queensland indicate predominantly reverse faulting. This faulting is the result of the crust being compressed horizontally in a NE-SW direction.Activity in the Burdekin Falls Dam area of central eastern Queensland is confined to the Drummond and Bowen Basins. There is also significant activity offshore from Townsville and Bowen.

Jan 1, 1990

By Paul Beck

IT is well known that the hardness of metallic single crystals, like that of polycrystalline metals, increases during deformation (hardening by cold-work). It is also known that, as a consequence of deformation, the metallic material acquires an evident thermodynamic instability, revealed, for example, by the fact that the increased hardness gradually decreases with time, especially at elevated temperatures (softening by annealing). The temperature at which the stabilizing process begins is different for various metals, and for any metal it depends on the amount of deformation. Such stabilization of deformed metals may proceed in two different ways: through recovery, and through recrystallization. The former process allows the unstable deformed metal body to approach stability gradually and without visible change in the grain structure; the latter process, like an allotropic transformation, yields directly the undeformed stable "phase" in the form of new undeformed crystals, which, starting out of a certain number of nuclei, grow into and' consume' the deformed hard material. The extent of stability attainable through recovery is limited, espe-cially if the deformation of the metal crystal or crystals involves not only the gliding along slip planes, but also the bending of the slip lamellae, or at any rate the bending of comparatively large parts of the crystal lattice. In such cases, for instance in bent single crystals or poly-crystalline metals subjected to any deformation, recovery will not as a rule be able to restore the "straight" undeformed lattice, and stability can be attained only by recrystallization; that is, by the consumption of the bent crystal lattice parts by new, undeformed crystals-a much more drastic process. It is evident that the part of the surplus free energy stored in deformed crystals, which is present in the form of the elastic energy of the bent lamellae, is in these instances not considerably involved

Jan 1, 1937

By Paul A. Beck

It is well known that the hardness of metallic single crystals, like that of polycrystalline metals, increases during deformation (hardening by cold-work). It is also known that, as a consequence of deformation. the metallic material acquires an evident thermodynamic instability, revealed, for example, by the fact that the increased hardness gradually decreases with time, especially at elevated temperatures (softening by annealing). The temperature at which the stabilizing process begins is different for various metals, and for any metal it depends on the amount of deformation. Such stabilization of deformed metals may proceed in two different ways: through recovery, and through recrystallization. The former process allows the unstable deformed metal body to approach stability gradually and without visible change in the grain structure; the latter process, like an allotropic transformation, yields directly the undeformed stable "phase" in the form of new undeformed crystals, which, starting out of a certain number of nuclei, grow into and " consume " the deformed hard material. The extent of stability attainable through recovery is limited, especially if the deformation of the metal crystal or crystals involves not only the gliding along slip planes, but also the bending of the slip lamellae, or at any rate the bending of comparatively large parts of the crystal lattice. In such cases, for instance in bent single crystals or polycrystalline metals subjected to any deformation, recovery will not as a rule be able to restore the "istraight" undeformed lattice, and stability can be attained only by recrystallization; that is, by the consumption of the bent crystal lattice parts by new, undeformed crystals—a much more drastic process. It is evident that the part of the surplus free energy stored in deformed crystals, which is present in the form of the elastic energy of the bent lamellae, is in these instances not considerably involved

Jan 1, 1937

Bakelite is fairly widely known as a suitable material for the mounting of ore specimens, concentrates and tailings for microscopical study, but has not as yet had extensive application. It is the most suitable mount for specimens which are to be held in the mechanical holder of a polishing machine, based on that described by J. W. Vanderbilt*, which is installed in the Geological Department of the Melbourne University.Bakelite products are prepared by subjecting bakelite powder to a temperature of about 1600 C. under a pressure of approximately two tons per square inch for a period of four minutes. The shape and size of the products are governed by the use of suitably constructed dies. When the concentrate or tailing is mixed with the bakelite powder, a briquette is produced which can be polished so that the opaque minerals in the sample can be examined microscopically in the same way as in a piece of solid ore. A simple machine for the construction of these briquettes has been described by Head and Slavin. It consists of a small four-pillar press in which the pressure is supplied by...

Jan 1, 1933

By Whincup P

Hilton deposit comprises en-echelon silver-lead-zinc orebodies with reserves in excess of 50 million tonnes. The host rock to the ore is a dolomitic and pyritic shale/siltstone which dips steeply west. The shale/siltstone is deeply leached and porous, particularly over the Ore Zone where the depth of leaching continues down to 440 m below surface, and constitutes an extensive permeable aquifer system containing a large volume of saline acidic groundwater. Initial production commenced at Hilton in 1987 well below the aquifer. A dedicated dewatering and monitoring programme is in place to ensure security of stoping operations and allow extraction of ore close to the leached zone in future years. Dewatering is undertaken from controlled bores from underground sites. Mine groundwater discharge is presently at the rate of 4000 kl/day with plans to increase to 8000 kl/day. An additional 2000 kl/day of fill and service water and water from secondary aquifers is also pumped from the mine daily. Although recharge rate to the aquifer is minimal, studies indicate that due to the size of the aquifer and future extensions to mine development, dewatering will be a life of mine activity. The investigations which have been undertaken to determine optimum dewatering techniques and schedules are described. These include production bore tests, mine planning constraints and computer aquifer modelling. Gravity geophysical techniques have been successful in delineating the lateral and vertical disposition of the aquifer. The mine dewatering schedule, whose principal component is underground mine drainage, is also outlined.

Jan 1, 1988

By V. A. Josendal, J. W. Wilson, B. B. Sandiford

Two radioactive tracers have been tested as a means of determining core saturation in multiphase flow studies. Cesium chloride was tried as a water-phase tracer, but complications in its use in low permeability cores resulted from sorption of cesium by the core or water-wet pads. IodoI3'ben-zene proved very satisfactory as an oil-phase tracer. The synthesis of iodobenzene from the sodium iodide as received from Oak Ridge is simple and direct. The tracer is insoluble in water and there was no evidence of sorption by any of the core materials used. use of the method to determine saturation profiles during capillary and dynamic desaturations and relative permeability measurements on oil-water and oil-gas systems is described. Comparisons of the dynamic and capillary methods of relative permeability determination were made using the tracer to check core saturation and saturation distribution. Other experiments are also described in which mobility of the oil phase at various saturations was measured by displacing labeled oil by flowing inactive oil. Similar experiments were made using water labeled with cesium 134. INTRODUCTION To be fundamentally sound. any laboratory method of determining relative permeability must meet, among others, the following requirements: (1) the core saturation between the pressure taps must be uniform, and (2) there must be uniform pressure difference between the phases in the region between the pressure taps. In practice it may eventually be shown that appreciable departure from either or both of these conditions may be permitted without introducing sensible error. but the magnitude of such error remains to be established. There is need then for methods of saturation determina- tion which will determine the saturation profile in a core as well as the overall average saturation given by the gravimetric or material balance methods usually employed. Methods which have received consideration include resistivity, X-ray absorption, gamma ray absorption, neutron diffraction, and radioactive tracers. The resistivity method has use in the determination of profile, but needs checking by independent means.' The X-ray method has been used successfully, but requires elaborate equipment and calibration.23 4 At present gamma ray absorption is practical only for very large cores. The neutron methodo does not appear adaptable to measurement of saturation variations in short cores since it is a scattering method. Russell, Morgan and Muskat' employed radiovanadium in a study of the mobility of interstitial water. Coomber and Tiratsook used radioiodine as an oil-phase tracer and measured profiles in unconsolidated sand packs. ADVANTAGES OF RADIOACTIVITY METHOD A thorough discussion of the theory of radioactive tracers and the technique of using them to study the movement of fluids in sands is given by Coomber and Tiratsoos and need not be detailed here. The method requires less equipment than the X-ray method; the counting problem is similar, but the generation and regulation of the radiation, which is a major problem in the X-ray method, is not required in the radioactivity method. In the X-ray method it is customary to add up to 20 per cent of absorber, usually an iodine compound, to the phase to be labeled. while in the radioactivity method a mere trace is sufficient. A possible disadvantage of the radioactivity method is that a disproportionate amount of the activity comes from the portion of the core nearest the counter, whereas the X-ray gives a true average saturation in the region traversed by the beam. 011 the other hand, this characteristic of the radioactivity method renders it peculiarly adaptable to study of certain types of radial variations in saturation. As with the X-ray method, a tracer may be added to either the oil or water-phase. An oil-phase tracer is more generally useful. however. since it permits work on oil and gas in the

Jan 1, 1952

By Hoyt Gale

RECENT work on borate deposits near Kramer, in the extreme southeast corner of Kern County, California, is of special interest to those who are making a study of the mode of origin of the borate minerals colemanite and ulexite. A valuable clue to the mode of origin of colemanite in its present known deposits is given in a paper by W. F. Foshag1, of the United States National Museum. Colemanite is found in or near original deposits of ulexite, from which it is supposed to have been derived by a change in the rate of movement or the character of the ground water. Indications that an alteration of this sort had produced colemanite were first recognized by the writer' in specimens collected in the Kramer district in 1920, and sent to W. T. Schaller, at U. S. Geological Survey, for identification. The identity of the two borate minerals in these specimens was not at once determined, but the relations of the minerals as seen in the specimens afforded a manifest clue to the origin of the deposits. Similar specimens were obtained by the writer from deposits in death Valley soon after he had examined the deposits near Kramer. After studying the specimens submitted to him Dr. Schaller wrote as follows: "I have finally positively identified the boron minerals you sent me with your letter of Sept. 27. The `white fibrous mineral like ulexite' is ulexite; the harder nodules' in the clay are colemanite, apparently formed from the ulexite as a result of leaching, with recrystallization in place-a theory that Foshag is working on." The present paper is a report on the facts observed at. the deposits near Kramer, and some general observations on the geological relations noted, which, unfortunately, still form only a very incomplete record.

Jan 2, 1926

W. A. SCHLESINGER,* Denver, Colo.-Two problems are of great interest to the radium manufacturer at the present tine: The first is to perfect a process, more efficient and more economical, which is capable of treating a greater variety of ores. . Practically all the radium now manufactured in the United States is extracted from carnotite ores and these vary a great deal according to their location. The amount of sulfates contained in the ore is of particular importance, and any straight acid-leaching method, such as developed by the Bureau of Mines, becomes prohibitive when the ore contains more than 0.5 per cent. of sulfates, as the sulfates go into solution and naturally precipitate the radium as insoluble radium sulfate. Since the amount of ore obtained from any one claim is usually comparatively small, it is necessary to work a number of claims in different localities, and a process, in order to be commercially satisfactory, should handle all of these ores. In other words, it would be entirely unprofitable for radium manufacturers to employ any process which cannot treat any but sulfate-free ores. He must deal with carnotite containing appreciable amounts of sulfates, and must be able to work this in with the other ores. Carbonaceous ores from certain sections in Utah; which are known to contain as pinch as 1.0 per cent. and more of volatile matters, are prone to a, nitric acid-leaching process. The second problem relates to the treatment of concentrates. There is no doubt that a very large amount of low-grade ore, containing less than 0.75 per cent. uranium oxide, is available. It has been demonstrated that these ores can be concentrated, but there is still room for research

Jan 11, 1918

By H. G. Doll

A new logging method, called induction logging, is described; it measures the conductivity, or resistivity, of the strata traversed by a bore hole. The apparatus, which is briefly described, comprises a coil system coupled with the ground by induction, so that there is no need for direct contact with the mud, or with the formations. For that reason, the method is particularly useful in oil base mud, and it was first applied to that case. A discussion concerning the respective contribution of each region of ground in the vicinity of the coil system is given. Through the concept of a geometrical factor, it is demonstrated that the total signal, which measures the apparent conductivity, is the summation of the partial products — geometrical factor by conductivity — for the different regions of the ground under consideration. It is thus possible to establish, for each type of coil arangement, investigation characteristic as well as departure curves and correction charts for beds of various thicknesses. This should greatly facilitate the quantative interpretation of the field logs, even in cases where such interpretation would have been difficult on electric logs taken with electrodes. It is also shown that focussing combinations of coils can be used, which give, directly on the recorded log, a value close to the true conductivity, even for fairly thin beds, thus eliminating in most cases the need for corrections. These focussing systems also give a sharper definition of the boundaries between successive strain.

Jan 1, 1949

By H. G. Doll

A new logging method, called induction logging, is described; it measures the conductivity, or resistivity, of the strata traversed by a bore hole. The apparatus, which is briefly described, comprises a coil system coupled with the ground by induction, so that there is no need for direct contact with the mud, or with the formations. For that reason, the method is particularly useful in oil base mud, and it was first applied to that case. A discussion concerning the respective contribution of each region of ground in the vicinity of the coil system is given. Through the concept of a geometrical factor, it is demonstrated that the total signal, which measures the apparent conductivity, is the summation of the partial products — geometrical factor by conductivity — for the different regions of the ground under consideration. It is thus possible to establish, for each type of coil arangement, investigation characteristic as well as departure curves and correction charts for beds of various thicknesses. This should greatly facilitate the quantative interpretation of the field logs, even in cases where such interpretation would have been difficult on electric logs taken with electrodes. It is also shown that focussing combinations of coils can be used, which give, directly on the recorded log, a value close to the true conductivity, even for fairly thin beds, thus eliminating in most cases the need for corrections. These focussing systems also give a sharper definition of the boundaries between successive strain.

Jan 1, 1949

By Y. L. Yao

A scale based on existing data for the mutual solid solubilities of metals in binary systems has been compiled. This scale is useful for estimating the relative magnitude of two solubilities, provided that these two metals do not form a continuous series of solid solutions. In his classic work on the theory of metallic solid solutions, Hume-Rothery proposed that, other things being equal, a metal of lower valency should be regarded as more likely to dissolve one of higher valency than vice versa. He called this the "relative valency effect". Later he found that the application of this general principle was limited to the three univalent metals: copper, silver, and gold. In dealing with alloys of these metals with metals of higher valency, Hume-Rothery stated, it is usually found that the solid solution of the univalent metal is of the greater extent.&apos; The question of whether there is a systematic trend in comparing the relative magnitude of mutual solubilities, if valency is not considered a criterion, naturally arises. The present paper proposes a "mutual solid solubility" scale on an empirical basis. The order of its arrangement is determined by the relative magnitude of mutual solid solubilities. Some applications of and inferences from this scale are given. ORIGIN AND SCOPE OF THE SCALE An extensive survey of over 400 binary phase diagrams shows that, with certain exceptions and within certain limitations, metals can be arranged in the following order: Na Zn Cd Ga Bi Ge Sb As Hg Sn Si Mg Li Ba In Tl Pb Al Ca Be Mn Cr Re W Mo Cu V Ta Fe Ru Cb Co Ti Hf Ni Os Zr Ir Pt Au U Rh Ce Pd Ag Th so that a metal higher on the scale tends to have a higher solubility in a metal lower on the scale than vice versa. some experimental solubility data are given in Table I. These data, unless otherwise specified, are taken from Hansen and Anderko.6 Throughout this paper, the term "solubility of A in B" is taken to mean the maximum solid solubility of A in B in atomic percent, while the maximum solid solubility of B in A in atomic percent is called "reciprocal solubility". The exceptions found in the above-mentioned survey are: Al-Ag system, 20.34 at. pct Al in Ag, 23.8 Ag in Al; Mg-Pb system, 5.9 Mg in Pb, 7.75 Pb in Mg; As-Sn system, vanishingly small amount of As in Sn, a large As primary solid solution; Cd-Hg system, Cd in Hg <Hg in Cd; Mg-T1 system, 5(?) Mg in ßT1, 15.4 T1 in Mg; W-Ti system, -21 W in ßTi, -25 Ti in W; Zr-Ag system, probably small amount of Zr in Ag, 20 Ag in ßZr.7 Considering the large number of examples, the exceptions are few. It is remarkable that metals may be so arranged without considering other factors such as size, crystal structure, electronic structure, and so forth. A limitation to the use of the scale is the fact that it cannot be tested for two metals forming a continuous series of solid solutions. It is interesting to point out that when the continuous solid solutions have a miscibility gap the critical composition is

Jan 1, 1961

An on-stream iron ore analyser has been developed in collaborationwith Hamersley Iron Pty. Limited for determining the iron content ofboth lump (- 30+ 6 mm particle size) and fines ( - 6 mm particle size)on conveyor belts. The analyser, which is called IRONSCAN, is basedon pair production, and is now commercially available from Mineral ControllnstrumentationLimited (MC!) in Adelaide. It can be mounted underexisting conveyor belts with minimal modifications to the conveyor structure,and the presence of steel cables in the belt does not interfere oncethe analyser has been correctly calibrated. The analyser has been extensivelytested on both lump and fines on the shiploading conveyor at Dampier,and typically the root mean square (Lm.s.) deviation between singleIRONSCAN measurements and conventional chemical analyses is betterthan 0.5070 Fe. It is currently being evaluated on - 150 mm ore from theprimary crusher at Mount Tom Price, and initial result are encouraging.The principal advantage of IRONSCAN is that it provides rapid informationon ore grades. But perhaps its greatest potential is at the primarycrusher where conventional sampling and analysis is very expensive toimplement.

Jan 1, 1988

By Mark Sheppard

DURING the winter of 1937, the writer visited a West Virginia stone quarry at which the overburden is stripped hydraulically. The quarry is in a bed of limestone, about 200 ft. thick, which outcrops on a hillside above a small river; the quarry face is 175 ft. high. The top of the stone is cut in all directions by erosion channels, which are filled with a tough, red clay. Some of the larger channels are 30 ft. deep and 8 ft. wide. The thickness of the clay overburden ranges from a few inches to 10 ft., the average being estimated as less than 3 feet. The topography of the district is rugged, and the hillside above the quarry face is not suited to the use of the usual mechanical methods of moving dirt. Until hydraulicking was started 10 years ago, the over- burden was dug by hand labor, dumped to the quarry floor and transported to the dump in quarry cars. Cleaning deep channels was a slow and expensive process. When hydraulicking was started, the top of the quarry face was over 200 ft. above the level of the river. The pipe line was carried in a semi- circle around the top of the quarry about 200 ft. from the face and about 50 ft. above it. As there is no adequate supply of water above the quarry, the water must be pumped from the river. As quarrying progressed the pipe line was moved until the discharge points are now over 300 ft. above the river. Water is pumped from the river by a four-stage centrifugal pump having a 5-in. intake and a 4-in. discharge. It is direct-connected to a 200-hp., 2300-volt, 1760-r.p.m. induction motor. At 1760 r.p.m. the rated capacity of the pump is 750 gal. per minute under a head of 600 ft. The pump is installed in a house of frame construction on the river bank, about 10 ft. above normal water level, which also houses the pump for the mill-water supply. Above the pumps are hoisting bays into which the motors are raised whenever the river reaches flood stage. An 8-in. suction line, equipped with a strainer and a foot valve, takes the water from a concrete sump at the edge of the river. The pipe line from the pump consists of 1750 ft. of 8-in. spiral steel flanged pipe and 800 ft. of 4-in. spiral steel flanged pipe. The connection from the pump

Jan 1, 1938

By Cam T. Nhan

The potential use of coal fly ash as a construction material has been recognized for many years,. In this work, an admixture of the ash residue is investigated for use as a landfill liner. In particular, its performance as a chemical barrier to the leachate from the municipal solid waste (MSW,) is determined. The test specimens are a water compacted admixture of class F fly ash, low in calcium content, which is blended with lime dust and calcium bentonite. MSW is passed through the compacted fly ash specimens under a gravitational head. The. diffusion of elements in the leachate through this material is monitored as a function of time. The distribution profiles of the retained elements in the compacted admixture provide prediction of the long term performance of the compacted admixture directly affect the leachate diffusion rate, hence, the mobilities of the retained elements.

Jan 1, 1992

By J. F. Wolff

EXPLORATION of Lake Superior iron ores is done principally by drilling. The soft iron ores are churn drilled and the harder ores are diamond drilled. In exploratory work in the harder formations, where the diamond drill is used, the sample comprises both core and cuttings. The care given in handling and preserving the core depends on the im-portance that may be attached to certain characteris-tic strata in different formations. The following instructions for the care of core samples are issued by the Oglebay Norton Co. for diamond drill work done for them on the Gogebic Range of Michigan: The driller shall be careful upon removing the core from the core-barrel to keep the core in exact order. Not only should the pieces of core be kept in exact relation to each other, but care should be used not to turn them end for end. Core should be placed in the core-box from left to right in each row. Each drill should be furnished with small, hinged cover core-boxes, and each drill should be well supplied with empty boxes. Core remaining at the drill when the drillers are not working should be locked up. The practice of picking sample pieces of core and carrying them away from the drill to be examined in better light, or shown to someone else, is unnecessary and prohibited. The reason for this is that these pieces are seldom returned to the core-boxes and valuable information or "markers" may be lost. If the inspector suspects that a critical point is reached in the drilling and he cannot determine at the drill, he should remove the entire core-box and see that the drill is immediately furnished with an empty core-box. It is customary to split the core longitudinally so that one-half may be ground to a pulp for analysis and the other half may be kept for future inspection. The core should be filed in permanent core-boxes, properly labelled, if the core is to be kept in exact order as taken from the core-barrel. Otherwise, it is often filed in tin boxes inn the way that sludge samples are filed. To obtain the average analysis, the core and cutting samples obtained in the same run are analyzed separately and the analyses are combined in the proper ratio.

Jan 9, 1922