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By J. B. Cheatham

The influence of friction on the force required for an idealized bit tooth to penetrate a "plastic" rock is analyzed. The rock is assumed to obey the Coulomb-Mobr yield criterion and the tooth is represented by a two-dimensional sharp wedge. At the tooth-rock interface the stress field not only satisfies the yield condition but also is such that the ratio of the shearing stress to the normal stress is equal to the coefficient of friction. The solution relates the force on the tooth to the depth of penetration, the tooth angle, the coefficient of friction between the rock and bit tooth and the angle of internal friction and compressive strength of the rock. INTRODUCTION Most rocks are brittle at atmospheric pressure but exhibit ductility and increased strength as the confining pressure is increased. In a previous analysisl the force required for a bit tooth to penetrate a rigid-plastic rock was computed for two limiting conditions: (1) a perfectly lubricated tooth-rock interface which should correspond to a lower limit for the penetrating force, and (2) a perfectly rough interface which should provide an upper limit. Experimental results2 for rock indentation under confining pressure have confirmed that the analysis generally predicts upper and lower limits for the penetrating force as a function of rock properties and tooth angle. The present analysis extends the previous work to account for the friction between the rock and the tooth. ANALYSIS The force required for incipient plastic flow of an idealized model of a rock under a bit tooth is computed subjected to the following assumptions. BIT TOOTH 1. sharp two-dimensional wedge (end effects neglected) 2. rigid tooth (elastic deformation of tooth neglected) 3. displacement of tooth is normal to rock surface 4. friction exists between tooth and rock ROCK 1. semi-infinite, isotropic and homogeneous 2. rigid-plastic (elastic deformation of rock neglected) 3. non-work hardening (horizontal stress-strain curve) 4. Coulomb yield condition* (linear Mohr envelope) 5. smooth, flat surface According to the Mohr theory of failure it is assumed that failure occurs when the state of stress is such that the Mohr stress circle becomes tangent to an experimentally determined envelope. The Mohr failure curve is frequent1y approximated by a straight line and referred to as the Coulomb-Mohr failure criterion. If the rock is under sufficient pressure to be ductile we can use the linear envelope as a yield condition. The Coulomb-Mohr yield criterion is illustrated by Fig. 1 and is expressed by the following equation:l

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Jan 1, 1942

Jan 1, 1897

Jan 1, 1914

By H. R. MacMichael

IN the Chihuahua district of Mexico the first smelting was that inaugurated by the early Spaniards for the production of silver bullion. The ores treated were high in silver and lead. Silver-lead bullion was first produced and the lead then cupelled to waste, with the slag produced running from 5 to 10 per cent in lead and about 25 to 40 per cent silica. The ore was mined by peon labor, often being carried out of the mine on the backs of the men, and then transported by burros. Furnaces were of primitive construction, adobe or rock being used in the earlier period and some iron later when it became available. Blast was originally supplied by bellows manually operated. Such smelting was carried on to a greater or lesser extent until the latter part of the last century, and so extensive was it that many hundred thousand tons of leady slag remained to be treated in modern furnaces. In 1908 a smelter was placed in operation at Chihuahua by the American Smelting & Refining Co. At that time the ore was practically all oxidized and by a little selection, self-fluxing mixtures could be made, but with an average lead content usually under 10 per cent and sometimes as low as 7 per cent. The bullion produced would carry from 150 to 300 oz. silver per ton. The slag made approximated 30 per cent SiO2, 42 per cent FeO, 21 per cent CaO, and perhaps 3 per cent Zn. The ores varied from hard, coarse material which was passed through a crusher, to a considerable quantity of fines which had to be smelted as it came, as crushing the coarse was the only preparation for smelting.

Jan 1, 1933

Jan 1, 1934

SUDBURY ore processed by International Nickel Co. of Canada Ltd. was recovered for many years as two mill concentrates, one primarily of copper, and the other mixed pentlandite-pyrrhotite, But, Since iron content of the nickeliferous pyrrhotite concentrate was ten times its nickel value, capacity of smelting was limited by the need for slagging the excess iron. A process for recovering the iron con- tent was developed after some years of research and a $20-million plant was constructed to employ the new method. Built near Copper Cliff, Ont., the plant started operation early in 1956, scheduled for separate treatment of 365,000 tons of pyrrhotite each year. It is only the first unit of an installation which will have a capacity of more than 1 million tons per year of pyrrhotite. By early 1958, production totaled more than 200,000 long tons of 68 pct Fe pellets containing 0.003 pct P, 1.5 pct SiO2, 0.01 pct S, and 0.15 pct Ni.

Jan 8, 1958

By F. R. Brotzen, J. Taranto

SEVERAL investigators have computed stacking-fault concentrations from X-ray diffraction data.'-' The method generally employed relates the line shift to the stacking-fault probability. In this investigation platinum filings of 99.99 pct purity screened to -325 mesh were studied after having been subjected to various annealing treatments. The lines were measured on a General Electric XRD-5 spectrogoniometer using CuK, radiation. The centroid of the broad reflections was determined by the method proposed by pike.' The stacking-fault probability was calculated by the formula given by Warren and warekois,3 with the assumption that all stacking faults were on the (111). Various high-angle reflections were used, but most data collected refer to the shift in the (400) reflections. When the filed platinum powder was annealed at 1200°C for 20 min the stacking fault concentration was too small to be detected. The stacking-fault probability in freshly filed specimens, as determined by the line shift relative to the annealed specimen, was about 0.016, Fig. 1. If the faults would be created only by extended dislocations, the dislocation density in the filed specimen would be about 3X1012 cm-2. The average stacking-fault width of about 21 which was used in this computation, is that given by Thornton and Hirsch.7 The mechanism by which stacking faults are annihilated during annealing is not yet fully understood. The isothermal results obtained do not seem to support the simple hypothesis of a first-order decay. On the other hand, the analysis by Kuhlmann, Masing, and Raffelsieper8 for the loss of dislocations during recovery agrees much better with the experimental data of this investigation. This analysis yields an activation energy of 29 ± 8 kcal per mol for the disappearance of stacking faults in cold-worked platinum during annealing. Even if one takes into account the lack of accuracy inherent in this X-ray method, the activation energy found is substantially less than that of self-diffusion in platinum, i.e., 61.6 kcal per mol.9 The results therefore suggest that processes other than dislocation climb, which involves the activation energy of self-diffusion, are responsible for the loss of stacking faults in platinum during annealing. The broad reflection lines became sharper early in the recovery process, Fig. 1. The sharpening occurred well before the stage during which stacking faults disappear. This lends support to Nowick's10

Jan 1, 1962

Jan 1, 1928

Jan 1, 1951

Jan 1, 1942

Jan 1, 1949

By Paul Hunt

UP to June, 1923, the Park-Utah mine had shipped about 94,000 tons of a direet-smelting ore of a gross value of $4,200,000, or about $45. a ton. These values were in gold and silver only, although the ore carried small amounts of lead, copper and zinc. Freight, treatment and deductions had averaged $15.40 a ton. At the expiration of the Pittman Silver Act there was in sight about 30,000 tons of a similar ore with a gross value of about $15 a ton in gold and silver, and a large body of partly developed lead-zinc ore of a gross value of about $20 a ton, 40 per cent. of which was represented by the lead-zinc content and the balance by the silver and gold. Mine costs had been high during the life of the Pittman Act, but as a large part of the profits had gone into equipment of the mine and the orebodies were large and had good walls, we knew that mine costs could be greatly reduced. We were therefore faced by the necessity of getting lower freight and smelter rates or of concentrating the ore. We did obtain better freight rates on lower grade ore, and a favorable smelting rate on the small tonnage of siliccous silver-gold ore, but this smelter rate could not be applied to the lead-zinc ore with any profit to ourselves. The Judge mill of the Park City Mining & Smelting Co., with which we were affiliated, was running to capacity; if we built our own mill it ineant a 5-mile freight haul to a satisfactory millsite and a year's delay. Fortunately, the International Sillelting Co. had just inherited a mill at Tooele which, with some changes, was about double the catpscity required for its own ores. Flotation tests were run on Park-Utah ore and a contract entered into with a low base treatment charge per ton, the tariff being largely fixed by the metallurgical deductions. This contract has been admirably calculated so that the miner can follow a lean orenbody to its limit. Such a procedure has no doubt made it difficult for the smelting company at times, but it has resulted in finding ore extensions that otherwise might have been missed, so In the long run it has profited too. A flat rate per ton would have been like mining up to the claim endlines in a great many of the stopes. In order to offer such a contract it can be inferred that the smelting company must have been able to make high reeoveries and high-grade products on a very lean ore.

Jan 1, 1928

Jan 1, 1910

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Jan 1, 1938

Jan 1, 1894

Jan 1, 1893

By S. A. Braley

DRAINAGE water from coal mines is probably the most serious water pollution problem today, varying in importance according to location of the mines and geological structure. Drainage may be either acid or alkaline in character. Acid discharge, the most severely detrimental to a stream, is caused by natural oxidation of the sulfuritic material (FeS2) in the strata associated with the coal seam. Since the acid is the result of a natural reaction the acid water differs because it does not cease with abandonment of the mining operations. There is no known economical method of neutralizing acid mine water or any practical method to prevent oxidation of exposed pyrite. Since production of acid from a mine does not stop when mining stops, the total quantity produced depends entirely upon the excavated areas. The increasing volume of acid water in manv mines has greatly increased operating costs. Pumping is expensive and acid mine waters are destructive of all equipment, especially metals in pumps and piping, and necessitate the use of corrosion-resistant materials. Discharge of the acid mine drainage into streams neutralizes their normal alkalinity, causes them to become acid, and produces an environment unfavorable for aquatic life and unsuited for industrial or domestic use without costly treatment. Mine drainages vary in percentage composition over wide limits, although the usual dissolved substances are ferrous and ferric iron, aluminum, calcium and magnesium sulfates, and lesser amounts of sodium, potassium and manganese sulfates and chlorides. Some alkaline discharges may contain heavy concentrations of iron as iron bicarbonate. These waters may produce iron hydroxide deposits in the receiving stream but do not cause it to become acid. In the extensive literature on acid mine water there appears to be a great deal of confusion about the importance of various components and the methods for their determination. In many instances faulty conclusions have been drawn from use of unsuitable methods of analysis. It is desirable that the factors and terms used in evaluation of analyses of mine waters should be so clearly defined that any interested person could properly appraise any analytical report. Some analysts report complete chemical analyses of mine waters and neglect to record drainage volumes. Others report only a minimum of analytical data after taking no precautions to preserve the original composition of the water during the time elapsing between collection and analysis. Some use methods of analysis intended for so-called pure waters of the potable and boiler water classes. These methods are not applicable to highly buffered waters such as mine water. Probably the most common criteria for evaluation are pH, free acidity, or acidity or alkalinity to methyl orange or methyl red, total acidity or acidity to hot phenol-phthalein, and the sulfate content. If these determinations are made on carelessly collected samples after a few days to weeks standing in warm rooms, they do not in any way represent the character of the water flowing from the mines. It is hoped that a brief discussion of the fundamental value of some of these factors may lead to a bettqr understanding of the need for more careful evaluation of mine water discharges. The term pH is one used by chemists to express relative acidity or alkalinity in terms of concentration of effective hydrogen ion in a solution. It is defined as the negative logarithm of the hydrogen ion concentration or activity in equivalents per liter. pH = logarithm A neutral solution, which is one containing the same number of hydrogen and hydroxyl ions, has a pH of 7. As the hydrogen ions increase and the solution becomes more acid, the pH decreases toward zero; and as the hydroxyl ions increase and the solution becomes more alkaline, the pH approaches 14. When dissolved in water to a dilute solution acids like sulfuric and hydrochloric, commonly known as strong acids, ionize completely, and the pH or hydrogen ion concentration varies with molar concentration of the dissolved acid. However, in high concentrations of such acids, the pH or hydrogen ion concentration is less than the acid concentration because the acid does not completely ionize. In only very dilute solutions does the pH represent the total amount of acid that can be neutralized by an alkali. All ionization reactions are equilibrium reactions. If other chemicals are added to the solution of an acid and the added chemical produces an ion that is the same as one of the ions of the acid, the degree of ionization of the acid is altered and the pH changes to some value that represents the active hydrogen ion of the new solution. Thus if iron sulfate is added to a solution of sulfuric acid the pH increases, since the common sulfate ion suppresses the degree of ionization of the sulfuric acid and decreases the effective hydrogen ion. However, the total acidity of the solution is increased. There are two salts composed of iron and sulfate— ferrous and ferric sulfate. In these salts there is no hydrogen ion that can ionize to give an acid solution, but when they are dissolved in water, the pH is less than 7 and the solution becomes acid. This is caused by a reaction known as hydrolysis and is represented by the equations FeSO4 + 2H2O ? Fe(OH)2 + 2H + SO,: or Fe,(SO,)3 + 6H2O ? 2Fe(OH)3 + 6H+ + 3504 A solution with a total acidity of 5000 ppm according to the first equation will have a pH of 4.40 but one with an acidity of 75 ppm, according to the sec-

Jan 1, 1958

By John Birkinbine

A specimen of brown hematite, taken from an iron-ore mine near Anniston, Alabama, exhibits a metallic streak or thread running throngh it. This specimen having been sent to me, I have no personal knowledge of its occurrence; but Mr. Albert E. Noble

Jan 1, 1895