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By W. Mitchell, B. H. Bergstorm, C. L. Sollenberger

A unique compression testing machine was constructed to load individual 1/8 to I-in. spheres of glass, etc., at rates from 100 to 100,000 lb per min. During loading the applied load was continuously plotted us the deformation of the sphere-platen system. The load-deformation curve conforms to the theory of elasticity. The spheres were surrounded by steel retaining rings or by gelatin (to prevent secondary fracture). The product size modulus was determined from screen analyses. The data shows: 1) Energy input per unit mass is inversely proportional to the product size modulus (or directly proportional to the surface area per unit mass). 2) Increased loading rates slightly increase the energy input per unit mass required for fracture without affecting the preceding relationship. Theory and experiments show that much of the input strain energy is transformed into kinetic energy of fragments at fracture, this being a reason efficiencies based on "surface energies" are exceedingly small. This large kinetic energy component is usually dissipated as heat but may be partially reclaimed because impact of fragments against the surface of steel retaining rings causes additional comminution. There are currently several theories of crushing, all of which can be derived from the general energy equation of Gillilandl as was pointed out by Mitchellet al in 1954'and more recently by Rose and Sullivan. If the exponent of the equation dE-x-"dx is assigned values of 2,1, or 3/2, one obtains by intergration between suitable limits, the equations of Rittinger, kick,' and Bond,' respectively. Other investigators prefer to let the exponent be a variable parameter. To help resolve the controversy, the Allis-Chal-mers Research Laboratories set up a series of carefully controlled experiments in which the level of energy input required to induce the comminution of a single particle by slow compression loading was very accurately measured. The energy input requirement, in conjunction with the size analysis of the fragments produced at fracture, yielded an accurate energy-product size relationship for the conditions studied. APPARATUS The experimental apparatus was designed to apply a compressive load on opposite sides of an "ideal particle." Spherical shapes were chosen, because point contact loading is typical of the mechanics of conventional crushers. The bulk of the experimental data was obtained with glass spheres because glass is a fairly homogeneous, brittle material for which the properties follow definite statistical distributions. The spheres ranged from 1/8 to 1-in. diam. The compressive load was applied to the specimen at a uniform rate by a loading mechanism (Fig. 1) which acted through a system of levers. The uniform rates, ranging from 100 to 100,000 lb per min, were obtained by metering water into a bucket suspended at the end of the secondary level. Point contact was made with the sphere through tungsten-carbide platens. This latter material is extremely hard and has a high modulus of elasticity, thereby having a minimum of distortion when loaded. The actual load on the sphere was measured independently of the lever system by a load cell, Fig. 2, which was designed so that the voltage output of the load cell would be proportional to load. The deformation of the specimen was indirectly measured by a pair of transformers which were attached to opposite edges of the upper platen. The cores of the transformers, which were attached to the lower platen, moved relative to the transformer so that the voltage output was proportional to the deformation of the sphere-platen system. The load was plotted against the deformation on an X-Y recorder. The area under the curve, as shown on Fig. 3, was proportional to the energy input of the

Jan 1, 1961

By Keith A. Heasley

At this point in time, approximately 800 M tonnes (880 M tons) of coal are mined each year in the United States. Forty percent of this coal or approximately 320 M tonnes (350 M tons) is taken from underground coal mines, and from this subtotal 16% or 50 M tonnes (55 M tons) is from longwall panels (1). If, for the sake of argument, an aver- age extraction thickness of 1.5 m (5 ft), an aver- age extraction ratio of 50 percent and an in situ density of 1360 kg/ms (85 lb/ft3) for this coal are assumed, then the 320 M tonnes (350 M tons) of underground coal translates to over 310 km2 (76,000 acres) that is undermined each year. Con- sidering that the total U.S. production of under- ground longwall coal to date is approximately 345 M tonnes (380 M tons), then a total of approximately 165 km2 (41,000 acres) (considering 100% extraction) has been directly undermined by longwall coal production in the United States. If the angle-of- draw is considered, the total area affected by longwall subsidence in the United States is probably twice the above value or 330 km2 (82,000 acres) . In the future, a number of other miscellaneous factors seem likely to multiply the number of potentially damaging situations that occur due to longwall coal mining. First, the overall production of coal in the United States is expected to gradually increase. This factor is further com- pounded by both the fact that the percentage of the overall production due to underground mining is bound to increase, since the vast majority of U.S. reserves (68%) are only recoverable by underground methods, and the fact that the percentage of the underground production of coal that is recovered by longwall mining is drastically increasing (1). Finally, as the nicest underground reserves are depleted, the remaining reserves will tend to be deeper, which increases the affected surface area, and under denser population concentrations, which increases the likelihood and severity of possible damages. Overall, considering the number of factor which are biased towards increasing the amount of undermined acreage, it appears that there must certainly be a corresponding increase in the number of possible damaging situations due to longwall-induced, ground subsidence.

Jan 1, 1986

By R. A. Frith

"Belts of Yellowknife Supergroup volcanic and sedimentary strata in the Slave Structural Province are separated by granitoid complexes. The strata and most intrusive rocks are 2.7 and 2.5 Ga old, but rocks around 3.1 Ga have been recognized in some of the complexes and are believed to represent a once-extensive sialic crust. Volcanics were erupted along rifts in this crust and the rifts became sites of crustal foundering and the development of troughs that filled with turbidites of volcanic and granitoid provenance. Present margins of the granitoid complexes may be close to the original margins of deposition of volcanics, with associated mineral deposits, and close to the thickest parts of basins of sedimentation.The Hackett River volcanic belt extends 100 kilometres between Mara River and Back River and is crudely S-shaped, with a northern sector surrounding a core of granitoid basement rocks, a northeasterly facin g central sector and a southerly facing southern sector. Up to 8 kilometres of andesitic volcanics with intercalated felsi c volcanic rocks are preserved within the central sector. The f elsic rocks include thick lenses of volcaniclastics with large, quartz-eye-bearing clasts, but the known metal deposits are not associated with these. Both of the deposit areas, the Hackett River area in the northern domal sector and the 'A itch ' Lake area 40 kilometres to the south in the central sector, are near the tops of relatively thin atypical volcanic sequences above suspected basement rocks. The Yava deposit near 'A itch , Lake overlies a unit of subaerially deposited, welded, andesit ic ash flows and the Hack ett River deposits overlie highly metamorphosed volcaniclastic rocks that include suspected ignimbrites.The deposits are notably argentijerous and lead-rich. They consist of stratigraphically zoned sheet-like bodies of iron sulphides, sphalerite, galena and chalcopyrite underlain by altered rocks containing chalcopyrite stringers. The Yava deposit was formed where hydrothermal fluids exuded through the fractured floor of a flooded caldera basin. The Hackett River deposits may have been formed in a similar setting and it is not inconceivable that all the known deposits in the belt were formed in a single large caldera basin, although there is no independent evidence for the required major fault separation between the Hackett River dome and the 'Aitch' Lake area."

Jan 1, 1980

By Hao Rang

The article Is specially to Introduce the recovery process of Nb, Mn, P from slags and reutilization techniques of the remaining slags. Nb-Mn-Fe alloys which contain 14-16% Nb, 40-60% Mn can be obtained after using fire refining process of " Blast Furnace-Revolving Furnace-Electric Furnace -Electric Furnace, "meanwhile, phosphorous slags which the content of phosphorus pentoxide Is 29.21%. Further going through wet metallurgical processing and electrowinning technics to recover grading 99°o metallic manganese. Finally, phosphorous slags are grinded finely to be as phosporous fertilizers to be directly used for farmlands. Small particle size (0-8mm) rich slags are mixed with slight solid additions for balling directly, and taken Into blast furnace with blocky slags all together. The remaining tall slags are fully used as constructional materials, slags cement as well as a serial of wall materials, road materials, ground ornamental materials without burning operation and steam curing. The compressive strength range of those materials is up to 15-30 MPa. There is no more solid waste slags effluent In the full technical process.

Jan 1, 1994

By K. A. Eriksson

Over 25 years ago, SKEGA pioneered the use of rubber linings in grinding mills and rubber is today widely accepted and almost standard material chosen for Regrind -and Secondary Mills. There are, however, only a few Primary Mills in operation with "solid" rubber linings because available materials and designs are often not cost efficient in comparison with steel alloys on a cost per ton basis and/or downtime for liner replacement would be too high.

Jan 1, 1988

Earl W. Baker W. P. Bigler Lyman Fearn Leslie H. Johnson V. O. Murray Max C. Scheble Fred L. Smith who left the world better for their having lived in it . . .

Jan 1, 1986

By C. S. Finney, John Mitchell

The decline of the beehive coking industry was inevitable, but it had filled the needs and economy of its day. A beehive plant required neither large capital investment to construct nor an elaborate and expensive organization to run. The ovens were built near mines from which large quantities of easily-won coking coal of excellent quality could be taken, and handling and preparation costs were thus at a minimum. The beehive process undoubtedly produced fine metallurgical coke, and low yields were considered to be the price that had to be paid for a superior product. Few could have foreseen that the time would come when lack of satisfactory coking coal would force most of the beehive plants in the Connellsville district, for example, to stay idle; and if there were those like Belden who cried out against the enormous waste which was leading to exhaustion of the country's best coking coals, there were many more to whom conservation was almost the negation of what has since become popularly known as the spirit of free enterprise. As for the recovery of such by-products as tar, light oil, and ammonia compounds, throughout much of the beehive era there was little economic incentive to move away from a tried and trusted carbonization method simply to produce materials for which no great market existed anyway. With the twentieth century came changes that were to bring an end to the predominance of beehive coking. Large new steel-producing corporations were formed whose operations were integrated to include not only the making and marketing of iron or steel but also the mining of coal and ore from their own properties, the quarrying of their own limestone and dolomite, and the production of coke at or near their blast furnaces. As the steel industry expanded so did the geographic center of production move westward. By 1893 it had moved from east-central to western Pennsylvania, and by 1923 was located to the north and center of Ohio. This western movement led, of course, to the utilization of the poorer quality coking coals of Illinois, Indiana and Ohio. These coals could not be carbonized to produce an acceptable metallurgical coke in the beehive oven, but could be so treated in the by-product oven. By World War I the technological and economic limitations of the beehive oven as a coke producer were being widely recognized. After the war the number of beehive ovens in existence dropped steadily to a low of 10,816 in 1938, in which year the industry produced only some 800,000 tons of coke out of a total US production of 32.5 million tons. The demands of the second World War led to the rehabilitation of many ovens which had not been used for years, and in 1941, for the first time since 1929, beehive ovens produced more than 10 pet of the country's total coke output. Production fell off again after 1945, but the war in Korea made it necessary once more to utilize all available carbonizing capacity so that by 1951 there were 20,458 ovens with an annual coke capacity of 13.9 million tons in existence. Since that time the iron and steel industry has expanded and modernized its by-product coking facilities, and by the end of 1958 only 64 pet of the 8682 beehive ovens still left were capable of being operated. Because beehive ovens are cheap and easy to build and can be closed down and started up with no great damage to brickwork or refractory, it is likely that they will always have a place, albeit a minor one, in the coking industry. The future role of the beehive oven would seem to be precisely that predicted forty years ago by R. S. McBride of the US Geological Survey. Writing with considerable prescience, McBride declared: "A by-product coke-oven plant requires an elaborate organization and a large investment per unit of coke produced per day. Operators of such plants cannot afford to close them down and start them up with every minor change in market conditions. It is not altogether a question whether beehive coke or by-product coke can be produced at a lower price at any particular time. Often by-product coke will be produced and sold at less than cost simply in order to maintain an organization and give some measure of financial return upon the large investment, which would otherwise

Jan 1, 1961

By Gordon I. McPhail, John D. Nelson

Tailings management represents a major portion of the cost associated with extraction of gold. For each gram of gold extracted, 150 to 200 kg of waste material is produced. Because of environmental concerns and unwillingness of society to accept environmental damage caused by poor waste management practices it has become essential to follow responsible and safe practices in tailings management. This paper presents general aspects of tailings management and discusses different techniques utilized. Characteristics of tailings and the effects that these have on stability and environmental aspects are discussed. The paper also discusses the effect of mine back-filling practices on the surface disposal of the remaining tailings fraction.

Jan 1, 1987

By Ibrahim AR

The experimental phosphate mine of abu Tartur has been recently placed into production. Some problems'have been encountered in maintai- ning grade control due to the differences between early block estimates and actual produ- ction grade. An area of 100 km2 is evaluated by using four methods of evaluation , namely polygonal, inverse distance , inverse distance squared and geostatistical methods. The results of calculations are compared statistically with the control values of grade and thickness. The comparison showed that the. kriging technique gave better estimations for the relative ore reserved, while for grades or thickness any of these techniques except the polygonal can be Used.

Jan 1, 1986

By A. D. Bauer

For several years the Bureau of Mines has conducted surveys to determine the changes in motor gasoline being sold throughout the United States. The present survey shows that far the districts in which samples were collected, the average gasoline is becoming more volatile instead of less so, as is sometimes supposed. This year's gasoline is much more volatile than that sold two years ago, and it has a somewhat better distillation range than last summer's samples. A comparison of the average figures for several years, shows that motor gasoline is also becoming more uniform in character. The large seasonal change is disappearing, but ?winter gasoline', still has a lower initial boiling point than "scanner gasoline". This difference in volatility is made intentionally to facilitate starting the motor in cold weather. The end point shown in the present survey is slightly lower either than that of last winter or of the summer of 1921.

Jan 1, 1922

By L. G. Closs

Gold continued to be the primary focus of minerals exploration worldwide during 1995. Other targets, though, included primary diamonds and base metals, especially copper deposits amenable to solvent-extraction/electrowinning (SX/EW) processing. The discovery of nickel, copper and cobalt at Voisey Bay in Labrador, Canada was an unusual surprise. Meanwhile, exploration geochemistry continues to evolve to meet new challenges. The major event in this category during 1995 was the17th International Geochemical Exploration Symposium (IGES) held in Townsville, Queensland, Australia in May. The meeting was sponsored by the Association of Exploration Geochemists, which celebrated its 25thanniversary of service to the mineral industry, and the Economic Geology Research Unit of James Cook University, Townsville.

Jan 1, 1996

By R. O. Williams

Analytical representation of the thermodynamics of solutions is highly desirable from the standpoint of accuracy, compactness, and numerical manipulations. In particular, computer calculations are greatly implemented. Mathematical considerations show that previous expressions have one or more serious defects. This investigation shows a Fourier series to be satisfactory but that it is also possible to derive a new series which fits certain additional conditions. Included examples show the value of analytical expressions in giving a simple characterization of each system using some two to five parameters, the elimination of the Gibbs-Duhem integration, and the es timation of the error for the experimental function as well as derived functions. It is further shown that the present characterization provides easy comparison between systems. IN the past, thermodynamic calculations have depended to a considerable extent on tabular and graphical methods. As the volume and precision of such data increase such methods become less satisfactory. Specifically, the selection of the optimum representation and the estimation of errors require statistical methods which in turn require analytical representation. The utilization of such data require further manipulations which are best done analytically for maximum precision. For example, phase equilibria are determined by common tangents to free-energy curves: a graphical determination is normally of low accuracy. As computers are increasingly used analytical representations become almost mandatory. Insufficient mathematical consideration has been given previously to the selection of empirical expressions. Those expressions having some theoretical justification are generally too inflexible and mathematically unattractive. We consider the problem in some detail and show that a Fourier series can be effectively used. Also a new series is defined which has certain advantages. ANALYSIS We wish to consider the analytical representation of the heat of mixing, AH, the excess free energy, ?Gxs, and the excess entropy, ?sXS, as a function of composition, X, for binary solutions relative to the pure components in the same state. When a distinction is not required, we use W to denote any one of the above functions. One may use a Taylor expansion around X = 0 to generate a power series. As the derivatives are un- known we represent the series as W = A + BX + CX2 + DX3 + EX4 + ... [l] where the constants A , B, C , ..- are to be selected to provide some optimum fit. For the extremes of composition W is necessarily zero so it follows that A = 0 [2a] B +C + D + E +••• = 0 [2b] Nonelectrolytes, which we are considering, appear to satisfy the condition that d3W/dx3 = 0 [3] in the terminal regions. This is the basis of the a, ß, and Q functions used by Hultgren et al.' and others. While this condition does not have a strong theoretical basis it does appear desirable that any analytical relation should satisfy this condition. Darken2 and Turk-dogan and Darken3 have shown that many systems exhibit this behavior over an extended range from each terminal region, departure being restricted to a limited intermediate region. Since we have no a priori knowledge as to where this transition occurs we can require that this condition be satisfied only as a limit at the extreme compositions as a general condition. We will show later how more restricted conditions can be included in specific solutions. Darken2 has called this behavior the quadratic formalism; we call our application the limiting quadratic formalism, LQF. This condition applied to the above power series requires that D = 0 [4a] 4-3-2E +5-4-3_F + 6 • 5 . 4G + ••• =0 [4b] The form of the power series normally used, due to Margules,4 is W=X(1-X)(A + BX + CX2 + DX3 + EX4 + •••) [5] where A, B, C, --. are a new set of constants. (Guggenheim5 has given a variation of this expression in a more desirable form. Since, however, it is contained in the above expression it does not require separate consideration.) This form is precisely what results by incorporating the conditions in Eq. [2] into the power series and regrouping the constants. The LQF requires that B =C [6a] and 4.3.2(D-C) +5-4-3(E-D) + ••• =0 [6b] Thus, the correct form of the Margules expression with two adjustable parameters is w =X(1-X)[A + B +X2-2/3x3)] 171 and the EX4 term must be included before three adjustable parameters are permitted.

Jan 1, 1970

By D. A. Whitley, R. P. Carone

"Shaft sinking by blind shaft drilling offers the mining industry a different approach to shaft construction. Benefits of blind drilling include less construction time and fewer safety problems while being economically competitive with conventional sinking methods under certain conditions in soft water-bearing formations.Technology and equipment are available today to drill blind shafts up to 20 feet in diameter and over 3,000 feet deep. This is an extension of proven techniques developed by private industry. To drill a 20-foot-diameter hole requires a specially manufactured hydraulic drill unit with a rated hoist of 2 million pounds and 500,000 ft/lb of torque.This brief paper will discuss the following aspects of blind shaft drilling:• benefits;• equipment f or drilling;• circulation methods;• circulation rates;• shaft alignment; and• progress rates to be expected.IntroductionShaft sinking by blind shaft drilling offers the mining industry an improved alternative method for shaft construction. The drilling of 20-foot -diameter shafts to depth s in excess of 3,000 feet can be done by utilizing equipment specifically designed and manufactured for this purpose. This equipment is a distant cousin, at best, to standard oil well drilling rigs. The actual configuration of these drilling machines will be discussed in a following section.To most people, the drilling of mine shafts may seem to be a new or at least recent technology. However, as early as the I920s, German and Dutch miners were drilling 24-foot diameter shafts for coal mines. What led ""miners to resort to drilling were ground conditions that still plague conventional shaft sinkers today: soft , water-bearing formations. The major drawback to these early attempts was a low penetration rate due to poor bit cleaning. This, however, was not the end of early shaft drilling activity. From 1936 to 1950, a number of mine shafts were drilled by the calyx or coring method. This method employs a bit with a large bore down the center. As it drills, it cuts a core. When 5 to 10 feet have been drilled, the core is removed. The disadvantages of this method were the time and work involved in breaking the core free, and it was inadequate in soft water-bearing ground."

Jan 1, 1981

By Walter C. Lawson

The primary objective in laying track in panel sections is to reduce the number of track laborers required. This is possible because the work is mechanized. Moreover, because the work is mechanized and each of the operations needs only a few men, and each is self-contained, the work can be carried on at night as well as in daylight; therefore, the method provides a means of preparing tracks on more than one shift. The laying of tracks in open pits and quarries in panel sections is not new but new methods have been made possible by the introduction of new types of equipment. The purpose of this paper is to describe the methods that are followed in the Morenci open pit. General Mining Operations Full-scale ore production at Morenci is about 50,000 tons daily. The normal yearly output of ore and waste is 32,000,000 tons, of which 30,000,000 tons is handled by rail haulage. This requires the building of approximately 47 miles of loading and other temporary tracks during a 12 months' period (Fig I). Broken ore and waste is loaded into 90-ton capacity dump cars with 5-cu-yd full-revolving electric shovels. Shovel banks are uniformly 50 ft high. They are blasted with churn drill holes and the broken material from a blast is loaded out with two shovel cuts, the first one being called (locally) the "splatter" cut and the second one the "clean-up" or "face-up" cut. When a shovel is making its clean-up cut, the loading track is about 70 ft from the toe of the bank and, under normal conditions, a bank is blasted against it without removing the track. It is thus evident that each track is used for two shovel cuts before relocation is necessary, inasmuch as a single position serves for both the clean-up cut prior to blasting and for the splatter cut following the blast. The normal advance of a solid bench by a single blast is 40 ft (Fig 2). Preparation OF Panel Grade In preparation of panel grades bulldozers and rooters are used for the bulk of the work and an auto-patrol road grader for the final stage. At Morenci, bulldozers have always been used for track-grade preparation, but before the introduction of the other two types of equipment grades were uneven and irregular at best, and always necessitated a big gang of track laborers to hand-block under the ties with rocks after a track was laid into place. High blocking also made poor track because of instability and caused frequent derailments. Under these conditions derailments were generally bad as re-railing became difficult. Much track was torn up and many tics broken in the process. The inadequacy of the hand-blocked track, together with the inefficiency of hand blocking, constantly pointed, to the desirability of reducing the amount of labor required, so various means were tried to improve conditions, such as wood

Jan 1, 1949

By B. L. Averba

THE determination of retained austenite by X-ray diffraction uses the following relationship:"" Pa = constant . RVaA (8) [1] where: P is the diffracted power from phase a; R, the calculated intensity factor; A, the absorption correction; 8, the diffraction angle; and Va, the irradiated volume of phase a For a flat sample making a glancing angle + with the incident beam the absorption correction is calculated as: * 1 sin (28 - 4) A .= 1 [2] µ sin, (28 — 4) + sin 4 where µ is the linear absorption coefficient. Beu3 has proposed that the computed absorption correction be combined with the constant and that the experimental observation of the absorption correction used in the original method be eliminated. This procedure is permissible only if: 1—There is no preferred orientation in the sample, and 2—The geometric requirements have been met precisely. If these conditions are not met the validity of the determination is questionable. Unfortunately the necessary conditions must be tested experimentally for each determination, and this is done most easily by observing whether the apparent absorption has the form of Eq. 2. In practice it may be convenient to plot P/R vs the parameter sin (2? — )[sin (28 — +) + sin +] to obtain straight lines for the ap- parent absorption correction. Deviations from straight lines indicate either that the sample has not been positioned correctly or that the relative intensities are incorrect. Geiger counter spectrometer methods also have been used successfully for retained austenite determinations. In the usual spectrometer arrangement the sample makes equal angles with the incident and diffracted beams (8 = ); thus the absorption correction is simply 1/2µ and independent of 8. The retained austenite is then easily calculated from integrated intensity measurements on austenite and martensite lines. If the austenite content is not too low and if excessive fluorescence is not encountered it is possible to use filtered radiation and a recording spectrometer. The integrated intensities are then proportional to the areas of the diffraction peaks. For lower austenite contents the following technique was used: A bent quartz monochromator and a Geiger counter spectrometer are employed. The monochromatic radiation is focused on the front slit, and the counter slit is made broad enough to include the entire line. A series of readings are taken in the vicinity of the diffraction peak. The highest reading includes the diffraction line plus background. The background is obtained from the average of several readings on each side of the peak. The integrated intensity is obtained by subtraction. CrK, and FeKa have been used for such determinations. This technique has been found to be as sensitive and as accurate as the photographic methods. In the spectrometer methods the ratio austenite/ martensite may be obtained from readings on austenite and martensite lines and the application of Eq. 1. It is also possible to use external standards and to obtain the austenite contents from measurements on the austenite lines alone. The small correction for carbides then becomes unnecessary. The best standards are probably samples with known

Jan 1, 1954

By H. M. Otte, A. L. Esquivel

A new leasl-squares method is Presented for determining lattice parameters of hexagonal or tetragonul structures. The method is adapted for use on electronic computers and involves a reiterative procedure. The correction factor employed raries linearly with the lattice parameter, a (determined from the Brag, angle). In contrast, Cohen's method and recent modifications of it use a correction factor that varies incersely as the squure of the lattice paramneter, a. While the recent modifications attempt to improve the precision of the extrapolated lattice parameter, a, (or-cu). by stressing the importance of the weighting factor. the present approach emphasizes the need for choosing the correct extrapolation function. A comparison between the present method (the Linear method) and Cohen's method indicates that the Linear method may be more appropriale in certain cases. through a priori no critertion appears to he available for making a chorce between the methods. The size of hexagonal and tetragonal crystal lattices is determined by two parameters, a, and c,. Although in principle the two lattice parameters can be determined independently from reflections for which h - k = 0 and 1 = 0, respectively, in practice this may be inconvenient (because of the angular positions at which these reflections may occur) or not easily possible (because of low intensity). Furthermore, if a high precision or accuracy is required, the limited number of reflections of this type available, particularly in the high-angle region, is not sufficient for the necessary corrections (mainly due to absorption) to be determined with accuracy. Several methods1-7,11-13 have been proposed employing all reflections, to obtain the optimum values, a. and co, either by a trial and error procedure or a least-squares fit. Of the latter method, cohen's5 is the best known one since it provides explicit expressions for the optimum a. and c,. However, Cohen's method is only strictly valid if an extrapolation function is used that varies linearly with l/a2 or lie2 (see Section 3), a requirement that does not appear to be generally appreciated. On the other hand, all the better known and more widely used A recent trial and error method was proposed by Massalski and King8,7 who computed extensive auxiliary tables of axial ratios vs the functions A = [(4/3)(h2 + hk +k2) + l2+(a/c)2] and C = A(c/a)2 used in computing a and c values from the observed Bragg angles. These values of a and c were then plotted against a function which permitted linear extrapolation. As a criteria; for the "optimum" values, Massalski and King rely upon a visual fitting of the line through points representing reflections of low 1-index points to compute the extrapolated value of ao and high 1-index reflections to obtain CO. The successive computations and graphical plotting required to reach the "optimum" value are quite lengthy and tedious even on a desk calculator and no quantitative assurance is obtained of having in fact selected the optimum value.* If the method of Massalski and King is used on an electronic computer, then their published tables become redundant and a least-squares fit becomes a natural selection for the choice of optimum values. Such an approach will be called the Linear method. For work now in progress on the effect of certain physical variables on the lattice parameters of hexagonal crystals, it has become essential to determine the confidence limits of small changes in the lattice parameters. Since extrapolation functions that varied linearly with a and c actually also appeared to vary linearly with l/a2 or l/c2 when tested against published as well as our data, a comparison of Cohen's method and the Linear method was considered desirable (Sections 3 and 4). For the latter method an electronic computer was required since a reiterative procedure to obtain the optimum ao and co values had to be employed. The purpose of this paper is to describe the principles of the Linear method, illustrate its application, and compare it with Cohen's method. 1) THE LINEAR METHOD The standard practice in obtaining the lattice parameters in the Debye-Scherrer method is to

Jan 1, 1965

By Raymond B. Ladoo

Cne of t·: . .:: i::,portt recent developrr.ents in the feldspar industcy l:.a.s beer. tl:.e r,:ipid. rise of Ne·N H.v-:psr.ire as a. feldspar producing State. Tl:.e unusual size and purity of t: e deposit now 'being worked ne3.r Keene, and tr.a i.nproved metl:.ods ot:' rr.ining srd ir,illing used l:.ave been responsible foe tt.is situation. ti'eldsp'.:l. w'ls occ':tsiona.lly produced sorr.e ye1rs ago in srr.a.ll (tuantities in New Hrunpshi :e, but t::.e State had not been considered a producer of" feldspar ui til lS3l. In the s::ring of tr .. at yea!', t:.e '[eane iic Products co. of Keene, N.H., beg3Il in a srr.9.ll way tl:e production of feldspar fro:n t1eir property in Gilsum, H.H., whicr. l-.a.d long been opera.tei a.s a :tl.ca mine. The purity of the feldspar at enoe '\ttracted f''.lvorai)le attention. and tt.e production of crude ore rapidly in­ cre'.\sed. Early in 1922, a new company, t'he Golding-Keene Co,

Sep 1, 1922

By Charles S. Barrett

THERE is a road into the microscopic realm that has remained untraveled through all these years of intense activity with high-power optical and electron microscopy. The road is worthy of careful scouting by metallurgists and other investigators, for it leads into relatively new country and it makes some areas accessible that heretofore could be reached only by very indirect paths or by the use of imagination. And when the road passes through familiar territory it frequently permits a new view of commonplace things that should prove of value from time to time. An accepted way to interest people in new sights such as this is to present an illustrated travelogue. This I propose to do, using many pictures taken in my recent exploration of this road. In the manner of travelogues, I shall touch briefly upon a large number of different subjects in physical metallurgy and shall, attempt to give an exhaustive treatment of none, for the objective is to provide a means for further research, to interest others in it, and to indicate what appears at the moment to be profitable directions for it to take. The method I shall present involves X-rays, but does not make use of the usual accompaniments of X-ray research such as reciprocal lattices, Fourier series, or the stereographic projection. The method is so simple in its fundamentals that it could have been developed 20 years ago, or perhaps even 30 years ago. It is well to have a destination for our travels, so that when we are through we shall have explored some region of importance, despite our preoccupation with the novel means of transportation and our interest in the sights to be seen along the way. So let us name the Cold Worked State as our destination, and make side trips into the territories known as Recovery, Recrystallization, and Precipitation-hardening. A great many investigations have been made of the nature of cold-worked metal, and a great variety of methods used, many of which are indirect, such as X-ray line intensities, X-ray line widths, calorimetric measurements, and magnetic measurements. Since comprehensive reviews of these investigations have been published in recent years, they need not be discussed here. The investigations that were the direct forerunners of the present work should be first mentioned. In 1931 I became interested in the structure of individual spots in X-ray diffraction patterns.1 When a crystal is placed in the narrow pencil of X-rays in a Laue camera, the spots recorded on the film are not always uniformly black, but may have an internal structure. The meaning of this may be understood from Fig. 1, which represents a narrow beam containing X-rays of many wave lengths ("general radiation") penetrating a crystal in the direction of the arrow and forming a pattern on the vertical film at the right.

Jan 1, 1945

By Walter E. Lewis

Ventilation-air-heating plants are in use at several iron mines on the Menominee Iron Range, Iron River, Mich. Many of the mines are exceptionally wet; during winter months, when fresh air is forced down the wet air shafts into the mine, it is usually necessary to heat the fresh air to above freezing so that the openings will not be closed by ice. The problem of ice closing air shafts has been solved by the practice of reversing the air current before the openings have become completely closed. This procedure involves drawing warm air from the mine up the usual down-draft air shaft. The M. A. Hanna Co. has found, at some of their mines on the Menominee Range, that it is more convenient to heat the below-freezing air rather than reverse the air current. Intermittent freezing and thawing of ice that has formed in the air shaft tends to loosen the wall rock, requiring major shaft repairs. The openings are kept free of ice the year around by heating the fresh air during periods of below-freezing weather. Table 1 shows the monthly minimum moan, maximum mean, and mean temperatures for the Iron River area. Under normal conditions of temperature, the furnaces are fired in November and usually are not allowed to go out until April of the following year.

Jan 1, 1951

The cut-and-fill mining methods in use in the mines of The International Nickel Company of Canada Limited, at Falconbridge Nickel Mines Limited, in the Sudbury District, and at McIntyre Porcupine Mines Limited in the Porcupine District, are typical of the general cut-and-fill practice in both the base-metal mines and the gold mines in the Province of Ontario. The following pages describe the methods in use. HORIZONTAL CUT-AND-FILL STOPING Horizontal cut-and-fill stoping, as practiced in the mines of The International Nickel Company of Canada Limited, falls into two main classifications: (I) transverse cut-and-fill stoping, where ground conditions or wide ore bodies make it necessary to establish narrow stopes and pillars across the width of the ore body; (2) longitudinal cut-and-fill stoping where ore body widths and ground conditions allow safe mining of the full width of the ore body. Transverse Cut-and-fill StopiNg General Stoping Layouts Stope and Pillar Widhs.—For transverse stoping, experience has indicated that narrow stopes and pillars, 27 ft. 6 in and 22 ft., respectively, are most satisfactory. These width are multiples of 5 ft. 6 in., which is the center to center dimension of a standard square set. Thus any stope, or block of stopes, can be readily converted from cut-and-fill to square set-and-fill if ground conditions should warrant such a change. Formerly, pillars were established 16 ft. 6 in. wide, the equivalent of the width of three square sets, with the intention of extracting the pillars by mining an overhand slice one set wide and an underhand slice two sets wide. However, subsequent experience proved that a two-set overhand slice is so much more efficiently mined than a one-set slice, and 22-ft. pillars have been made standard. Chutes.—The stope chutes are established on the longitudinal center line of the stopes and are spaced at 22-ft. centers from footwall to hanging wall. The footwall chute is usually from 5 to 10 ft. from the footwall of the stope. It is offset where necessary at an angle of 52° and vertical chutes are carried from the offset at 22-ft. centers. In sections of the ore body where a flat hanging wall would cut off hanging-wall chutes within a few cuts of the sill, it has proved to be more economical to use temporary chutes or slushers, or both, for these first few cuts, rather than to el permanent chutes at the hanging wall.

Jan 1, 1946