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By Fred C. Bond

MOST investigators are aware of the present unsatisfactory investigatorsstate of information concerning the fundamentals of crushing and grinding. Considerable scattered empirical data exist, which andare useful for predicting machine performance and give acceptable accuracy when the installations and materials compared are quite similar. However, there is no widely accepted unifying principle or theory that can explain satisfactorily the actual energy input necessary canexplain commercial installations, or can greatly extend the range of empirical comparisons. Two mutually contradictory theories have long existed in the literature, the Rittinger and Kick. They were derived from different viewpoints and logically lead to different results. The Rittinger theory is the older and more widely accepted.'TheRittinger In its first form, as stated by P. R. Ritted.'tinger, it postulates that the useful work done in crushing and grinding is directly proportional to the new surface area produced and hence inversely proportional to the product diameter. In its second form it has been amplified and enlarged to include the concept of surface energy; in this form it was precisely stated by A. M. Gaudin' as follows: "The efficiency of a comminution operation is the ratio of the surface energy produced to the kinetic energy expended." According to the theory in its second form, measurements of the surface areas of the feed and product and determinations of the surface energy per unit of new surface area produced give the useful work accomplished. Computations using the best values of surface energy obtainable indicate that perhaps 99 pct of the work input in crushing and grinding is wasted. However, no method of comminution has yet been devised which results in a reasonably high mechanical efficiency under this definition. Laboratory tests have been reported- hat support the theory in its first form by indicating that the new surface produced in different grinds is proportional to the work input. However, most of these tests employ an unnatural feed consisting either of screened particles of one sieve size or a scalped feed which has had the fines removed. In these cases the proportion of work done on the finer product particles is greatly increased and distorted beyond that to be expected with a normal feed containing the natural fines. Tests on pure crystallized quartz are likely to be misleading, since it does not follow the regular breakage pattern of most materials but is regularrelativelybreakage harder to grind patternat the finer sizes, as will be shown later. This theory appears to be indefensible mathematically, since work is the product of force multiplied by distance, and the distance factor (particle deformation before breakage) is ignored. The Kick theory4 is based primarily upon the stress-strain diagram of cubes under compression, or the deformation factor. It states that the work required is proportional to the reduction in volume of the particles concerned. Where F represents the diameter of the feed particles and P is the diameter of the product particles, the reduction ratio Rr is F/P, and according to Kick the work input required for reduction to different sizes is proportional to log Rr /log 2." The Kick theory is mathematically more tenable than the Rittinger when cubes under compression are considered, but it obviously fails to assign a sufficient proportion of the total work in reduction to the production of fine particles. According to the Rittinger theory as demonstrated by the theoretical breakage of cubes the new surface produced, and consequently the useful work input, is proportional to Rr-l.V f a given reduction takes place in two or more stages, the overall reduction ratio is the product of the Rr values for each stage, and the sum of the work accomplished in all stages is proportional to the sum of each Rr-1 value multiplied by the relative surface area before each reduction stage. It appears that neither the Rittinger theory, which is concerned only with surface, nor the Kick theory, which is concerned only with volume, can be completely correct. Crushing and grinding are concerned both with surface and volume; the absorption of evenly applied stresses is proportional to the volume concerned, but breakage starts with a crack tip, usually on the surface, and the concentration of stresses on the surface motivates the formation of the crack tips. The evaluation of grinding results in terms of surface tons per kw-hr, based upon screen analysis, involves an assumption of the surface area of the subsieve product, which may cause important errors. The evaluation in terms of kw-hr per net ton of —200 mesh produced often leads to erroneous results when grinds of appreciably different fineness are compared, since the amount of —200 mesh material produced varies with the size distribution characteristics of the feed. This paper is concerned primarily with the development, proof, and application of a new Third Theory, which should eliminate the objections to the two old theories and serve as a practical unifying principle for comminution in all size ranges. Both of the old theories have been remarkably barren of practical results when applied to actual crushing and grinding installations. The need for a new satisfactory theory is more acute than those not directly concerned with crushing and grinding calculations can realize. In developing a new theory it is first necessary to re-examine critically the assumptions underlying

Jan 1, 1953

By Fred C. Bond

DISCUSSION Milton C. Shaw and Donald R. Walker (Massachusetts Institute of Technology)—For many years the two empirical rules of comminution bearing the names of Rittinger and Kick have been applied to grinding processes to account for the energy required to crush brittle materials. The basis for Rittinger's rule was the thought that the energy to grind a substance went largely into supplying the additional energy associated with the new free surface, while Kick's law was prompted by the concept of dimensional similarity and is a special statement of the concept of constancy of strain energy to cause rupture. According to Rittinger the energy to grind a given volume varies as the change in the square of the mean particle size, v - A (D2), while according to Kick this energy varies

Jan 1, 1953

By G. G. Gunther, C. L. Boeke

THE zinc deleading processes at the Parral and Santa Barbara mills are described separately to provide a basis for comparison. Although the two procedures are fundamentally alike, there are some differences in application to meet specific conditions. The Asarco mill at Parral in the State of Chihuahua is a flotation plant treating 1450 metric tons per day of complex sulphide copper-lead-zinc ores. In general, the procedure used for selective separation of the minerals into a copper, a lead, and a zinc concentrate follows the usual practice. First, a bulk copper-lead concentrate is floated and then separated into a final copper and a final lead concentrate by the sulphurous acid process. A small tonnage of high grade gravity lead concentrate is tabled from the bulk copper-lead cleaner tailing before it is reground so as to avoid undesirable over-grinding and sliming of this portion of the lead mineral. Next, a zinc concentrate is floated from the lead rougher tailing and, as described later, is separated into a final lead product and a final zinc concentrate. Finally, the zinc rougher tailing receives a strong scavenger float to remove all recoverable middling values for regrinding and return to the flotation circuits. The slime is separated from the scavenger flotation product, and only the sand portion goes to the regrind section. The minerals in the separated slime are extremely resistant to all methods of selective separation and pose one of the major problems awaiting solution. Nevertheless, the slime shows a profit when mixed with either the lead or the zinc concentrate, except when metal prices are very low, so it is accepted as a final product for lack of a more profitable method of handling it. In contrast, the minerals in the reground sands of the

Jan 1, 1953

By G. G. Gunther, C. L. Boeke

DISCUSSION I. M. Symonds, (Cia. Minera de Penoles, Monterrey, N. L., Mexico)—Messrs. Boeke and Gunther have written most interesting papers describing their zinc de-leading operations. Mr. Gunther was kind enough to show me his operation last fall. Aside from the excellent results, I was most impressed by the ease of control. I had expected a difficult operation requiring careful control and supervision, but found one requiring little attention. The only disappointing fact of the separation is the poor copper recovery in the concentrate of the delead-ing float. The final zinc concentrate still has about 3 pct chalcopyrite, assuming the copper is chalcopyrite, and adversely affects the zinc grade. If the copper mineral is free, the poor recovery is probably due to the high cyanide concentration. Fortunately, about 75 pct of the copper is recovered before the pulp goes to the zinc circuit. Mr. Gunther's remarks about proper cyanide to zinc sulphate ratios cannot be overemphasized. Although our experience has been confined mostly to the conventional lead-zinc separation, we have found these ratios most important, and more important in the mill than in the laboratory. The laboratory appears to give optimum results over a wider range of reagent control than the mill, especially in regard to the effect of cyanide. In handling relatively high grade ores in the mill we usually find that ratios of cyanide to zinc sulphate of 1 to 10 or 20 or higher work best. Where there is a substantial amount of chalcopyrite present it has been noted that cyanide in the range of 0.02 to 0.10 lb per ton of ore is often beneficial in zinc rejection and possibly in copper recovery in the lead concentrate, while larger amounts, say 0.20 to 0.25 lb per ton, will float more zinc and cause a serious rejection of chalcopyrite. On an ore with 15 to 20 pct zinc content we have had to use 0.8 to 1.2 pounds of zinc sulphate per ton of ore with the above cyanide to obtain optimum results. Some time ago we made a laboratory study following Parral and Santa Barbara practice on an ore which had never a very satisfactory rejection of zinc in the lead concentrate. The results were not encouraging and a mill test was not run. Since then the study has been reopened using in the laboratory fresh zinc concentrate from the mill instead of one made in the laboratory. A reasonably clean lead float from the zinc concentrate was obtained. It has been checked several times and each time the zinc concentrate made in the mill responds and the one made in the labora-

Jan 1, 1953

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By Bunting S. Crocker

DISCUSSION Fred C. Bond (Allis-Chalmers Manufacturing Co., Milwaukee)—This is a very complete description of the screened ore fine grinding installation at Lake Shore. The only important data which might be desired are the operating screen analyses and power consumption in each stage. The statement that the energy consumption, or kw-hr per ton, remains constant for all mill diameters is of great interest, but the opinion is widely held that the kw-hr per ton decreases as the mill diameter increases. More definite test work is needed to establish this point. Screened ore fine minding has been successful on hard Lake Shore ore &d on about equally hard South African Rand ore. However, it is still not determined whether it would be successful on ores of average or less than average hardness. It is possible that with softer ores the consumption of ore pebbles might be exceedingly high, or might even exceed the available supply. Also, the softer feed might result in a rapid buildup of harmful intermediate sizes in the mill circuit. These factors can be determined only by a full scale or long pilot mill test. In any case the Lake Shore staff is to be congratulated for this pioneering work. Bunting. S. Crocker (author's reply)—As requested, the operating screen analyses and Bower consumption for each of the four stages of pebble milling are given, see Table XII. The fist three stages are in series circuit; the last stage is in closed circuit. The primary circuit had rock feed averaging 124 g per piece and the final stages rock feed weighing 63 g. The intermediate stages had mixtures of the two. Regarding Mr. Bond's statement that "the opinion is widely held that the kw-hr per ton decreases as the mill diameter increases," I would say that I have frequently been told this but have never yet been shown any evidence or seen any evidence published that would prove this statement. It may be true that some of the big mills recently installed have mechanical power savers such as micarta bearings and better lubrication, but as far as the power consumed inside the mill is concerned, we at Lake Shore think that it varies as D2 .", and that is all. We published extensive information on this point in our 1939 paper. Recently Mr. Banks at Consolidated published the figures on his 10 ft 8 in. x 12 ft 0 in. rod mill. As reported in this discussion by the writer: we calculated the power necessary to run this mill based on our 7x6-ft rod mill to within 1 pct of the actual power reported by Mr.

Jan 1, 1953

By F. E Turton

THE town of Rancagua, at the junction of the state and Braden Copper Co. railroads, is located 82 km south of Santiago, the capital of Chile. From Rancagua 70 km to the east, situated in the Andes at an elevation of 7000 ft, is the town of Sewell, the location of a millsite. Two and one half km east of Sewell by underground railroad at an elevation of 7500 to 10,000 ft is the Teniente ore-body. The history of El Teniente dates from the eighteenth century, when according to legend a Spanish lieutenant, or teniente, a fugitive from justice, escaped to the Andes and camped at what afterwards became known as the Mineral El Teniente. This man reputedly discovered a vein of high grade copper ore. The first official records date from 1819, when some exploitation was begun, and continue until 1870. The orebody was large and of relatively low grade and was beyond the capabilities of the small groups attempting to operate it. No effort was made to erect a concentrator until William Braden became interested in the property. In 1901 Mr. Braden promoted the necessary capital for the first practical exploitation of the ore-body. Barton Sewell and E. W. Nash, both of the

Jan 1, 1953

By Fredrick C. Pearson, Albert Brown, Emil R. Ermert

EARLY in 1946 the Shen-Penn Production CO., a subsidiary of the Philadelphia and Reading Coal and Iron Co., was organized to operate the Shenan-doah Stripping, one of the largest open-pit anthracite mines in the world. This mine is located in the Western Middle Field near Shenandoah, Schuylkill County, Pa., and involves the removal of approximately 60,400,000 cu yd of overburden to permit the recovery of 9,200,000 gross tons of rough cleaned coal from an area 8000 ft long and 2700 ft wide, with a maximum depth of 400 ft. Operations began in August 1946, and within the next few years 31,811,-000 cu yd of overburden and 2,994,000 gross tons of rough cleaned coal had been removed. During the early days of open-pit mining operations, steam was the primary motive power. However, as the size of the equipment increased, so did the problems of fuel, water, labor, and maintenance. Because of its efficiency and convenience, electric power rapidly replaced steam. Later the modern diesel engine was developed, and today practically all stripping equipment is either electric, diesel-electric, or diesel-powered. In locations where electric power is not readily available or where operations are scattered over a large area, diesel power is favored. At Shenandoah operations are highly concentrated in a relatively small area, and this fact, coupled with the availability of ample electric energy, indicated the selection of electric-powered equipment. Since this was a new company, it was necessary to purchase every item, from the smallest tools to stripping shovels. Although the organization period was hampered by the postwar scarcity of materials, an excellent opportunity was afforded to set up a thoroughly modern plant. This paper will describe the power facilities, with particular emphasis on the more unusual features. The substation is a major item in a power distribution system, and a simple arrangement utilizing high quality equipment is an excellent investment. With this in mind, a 3000-kva unit type substation was selected and installed. This unit is made up of three sections, which are assembled simply by bolting together the various sections and connecting the power leads, see Fig. 1. The transformer proper, rated at 3000 kva, is oil-filled and nitrogen-sealed, having a 66,000-v Delta primary and 4160-v Wye secondary with a grounded neutral and 46-ohm ground resistor. The function of this resistor will be described later. Electrical energy is purchased from a public utility and is delivered to the main substation at 66,000 v over one or the other of two incoming lines. These lines terminate on a high voltage switching section on which high voltage fuses and Thyrite lightning arresters are installed. The outgoing feeder switching unit is made up of the following: 1—An auxiliary compartment consisting of potential and current transformers with necessary terminal blocks for connecting to power company metering equipment. 2—Another auxiliary compartment containing a control transformer, battery, and battery charger for the furnishing of power for tripping and reclosing of power circuit breakers. The assemblies in these auxiliary compartments are mounted on carriages that can be rolled in and out with ease. Potential transformers are readily accessible for inspection and fuse replacement. When withdrawn from the operating position they are completely de-energized and grounded. 3—Two feeder breaker compartments, each with a 1200-amp, 5000-v, solenoid-operated magneblast air circuit breaker, voltmeter, ammeter, overcurrent relays, and a ground relay. These power circuit breakers can be quickly and safely withdrawn or repaired. A built-in isolating mechanism is used for lowering the breaker to a transfer truck for repair or inspection or as a means of disconnecting the breaker from the source of power as an added precaution when working on power lines. Thus inspec-

Jan 1, 1953

By W. H. Triplett

Silver-lead-zinc replacement deposits occur in limestones of the anticlinal Caja range at Providencia, near Avalos, Zacatecas, Mexico. The description deals with intrusives, outcrops, preferred horizons, changes from bedding to fracture, control in depth, and variations in abundance and ratios of minerals throughout a vertical range of 500 m. THE purpose of this paper is to record a few field observations and accumulated office data concerning outcrops, relation of ore occurrences to intrusive and host rock, and mineral zoning. Reasons for some of the phenomena are not clear to the author so the facts are given with the hope that other investigators may be able to furnish the explanations. The Avalos unit of the Compania Minera de Penoles, S. A., operates the Alicante, Bonanza, Providencia, Albarradon, San Eligio, Nazareno, Leona, Salaverna, and Santiago mines situated between Avalos railroad station and Concepcion del Oro on the Coahuila and Zacatecas railroad in the State of Zacatecas, Mexico. Some of these mines are within Penoles property while others are on mining claims leased from the Compania Nazareno y Catasillas, S. A., Mazapil Copper Co., and the American Smelting and Refining Co. The main office and camp are at Providencia, about 170 km southwest of Monterrey, N. L., so the silver-lead-zinc mines of the district are usually referred to as the Providencia mines. The Aranzazu, Catarroyo, and Cabrestante copper-gold mines are situated a few kilometers to the southeast on the same mountain range near Con- cepcion del Oro, the name by which the camp is known. It is operated by the Mazapil Copper Co. The main theme in the following pages is a description of the silver-lead-zinc ore occurrences of the district. Rock Formations Sedimentary Rocks: Many competent geologists have contributed generously to an ample fund of information concerning the sedimentary columns of northern Mexico. Among them are Carlos Burck-hardt, E. Bose, L. B. Kellum, and R. W. Imlay. Carlos Burckhardt, who covered a number of extensive areas in many parts of the Mexican Republic, devoted years to mapping the geology of the Caja, Zuloaga, and Santa Rosa mountain ranges around Mazapil, Concepcion del Oro, and San Pedro de Ocampo, Zacatecas. He did most of the original mapping which has been used as a basis and amplified by others. C. E. Burbridge and, more recently, John

Jan 1, 1953

By W. H. Triplett

DISCUSSION John G. Barry (Mexico)—"The Nerinea limestone," writes W. H. Triplett, p. 584, "is ... medium-bedded (0.5 to 1.0 m)...." This is the Zuloaga formation. It has always been described as thick-bedded (1.0 to 3.0 m). "Whether the marls and shales were less amenable to replacement," he goes on to say, ". . . is a matter of opinion." In general, it is probable that the thinner-bedded formations were less competent than the nearby thicker-bedded, so that the latter produced pre-mineral breccias when all were subjected to flexing stresses. With reference to Mr. Triplett's analysis, p. 584, under Lower Cretaceous, in the Taraises and Cupido formations some few horizons should be considered as possibly containing hidden orebodies. In the Potrero camp, shown on the general map in the northwestern part of Fig. 4, p. 587, several mineralized beds are known; they are found throughout the whole thickness of the Cuesta del Cura formation. With regard to intrusive rocks, p. 585, the general map, Fig. 4, shows that the Potrero and Noche Buena camps are near a third, monzonitic intrusive. It seems doubtful that the Aranzazu was any hotter than the Providencia camp. However, the former does contain large deposits of exogenetic silicates, which in large part occur within the Zuloaga limestone, as shown by Bergeats in a stereogram. At the Aranzazti camp the La Caja formation was only slightly changed, whereas it was altered highly at the Providencia, possibly en-dogenetically. That the temperatures at the Providencia camp were relatively high at one time is also shown by the tourmaline veinlets and lenses in the monzonite and by the garnets on Temeroso peak; high temperature mineralization has possibly been eroded. Maps and sections show that the Concepcion del Oro —Temeroso stock occupies the original locus of some 40 kms of sedimentary rocks. The remains of large roof pendants at Carmen and Animas indicate that stoping was active. Whether all the sedimentaries were stoped and assimilated presents a lively problem, a correlative of which would be an estimate of the original constitution of the invading instrusive rock. As to extrusive rocks, see p. 585, there is also a large thick area of the extrusives left near Noche Buena in the "La Caja" intramontane valley of the La Caja range. This place gave its name to the range. The volcanics are also said to occur in considerable amount in the Mazapil valley.'' Turning now to the remarks under Regional Structure, p. 586, as partly indicated on the general map, Fig. 4, the Mazapil Mountains of the Mazapil region are a continuous, U-shaped range including a northern leg of some 24 km, a southern leg of similar length, and a uniting base which lies to the East and which measures some 10 km in a S 20 E direction. Thus the U-shaped mountains are continuous through a length of 58 km and also contain one continuous anticline, the adjacent valleys being synclinal. The northern leg of the U is called the La Caja range, the base is called the Concepcion del Oro range, and the southern leg the Santa Rosa range. The ranges in northeastern Zacatecas and southern Coahuila contain anticlines and the valleys contain synclines, which where thin are necessarily highly compressed isoclines with very steep dips. Many of the anticlines also have isoclinal components which are overturned, original mushroom tops being indicated. The anticline in the base of the U of the Mazapil Region, Concepcion del Oro range, is a recumbent anticline, over to the ENE, with a necessarily subjacent recumbent syncline.

Jan 1, 1953

By C. L. Thornburg

FLANKING the Valley of Mexico on the northeast is a mountain range known as the Sierra de Pachuca. This northwesterly-trending range is about 30 miles long and 5 miles wide, its summit attaining an elevation of more than 10,000 feet above sea level, or 2000 feet above the valley floor. Pachuca, a town of 50,000 inhabitants, lies nestled at the southwest base of the range, 60 miles northeast of Mexico City. Three miles to the east, just over the summit and on the northeasterly slope, is the mountain town of Real del Monte with a population of about 20,000. These are the two principal towns in a 40-sq mile area whose yield has brought the district to its high rank among the world's silver producers. Total production probably exceeds 1.25 billion ounces of silver and 4.5 million ounces of gold. Exploitation by the Spanish was under way by 1530. As the mines were deepened operations became handicapped by the inadequate method of handling water with horse whims and bull skins. To offset this disadvantage a two-mile drainage tunnel was started in 1749 and completed ten years later. John Taylor of London acquired the mines from the third Count of Regla in 1824 and formed the Compania de Real del Monte y Pachuca. Two years later Cornish pumps were brought from England. A Mexican company purchased the British interest in 1848 and within a few years resumed an old project of driving a second, lower three-mile drainage tunnel, completing it in 1857. After Americzn interests acquired control of the property in 1906 operations expanded to an unprecedented scale, despite periods of political strife, unstable prices, labor problems, and a growing burden of taxation. An idea of the scale of operations may be gained from records of the last two decades prior to 1947, the last year for which production information is at hand, during which time production ranged from 1.4 million to 1.1 million tons per year, and the combined development and exploration, exclusive of diamond drilling, amounted to about 18 miles annually. Pachuca differs from Zacatecas and some other Mexican silver vein districts that flourished and de- clined in earlier centuries in that a substantial part of its total production was made after 1900. While the district's accelerated activity in this century was largely due to the successful application of the cyanide process to the treatment of the ores, improved pumping methods, and new mechanical equipment, it was at the same time stimulated by the opening up of important new orebodies as a result of active and extensive exploration of the vein system. Some of these orebodies were found on veins which terminate upward far below the surface, and though these particular veins do not crop out, they show relationship with a type of alteration which may be their surface expression. The district has been studied by many able geologists, and a large store of information has been accumulated, much of which is recorded in published as well as private papers. Wall Rocks Extrusive Rocks: The Sierra de Pachuca is a thick accumulation of Tertiary eruptive rocks consisting mainly of flows, breccias, and tuffs. Numerous recognizable volcanic vents within the range itself manifest that the range was built up by material ejected from these vents. The maximum thickness of the volcanic pile is not known, but from deep exploration and from relationship with Mesozoic sediments to the north and east it is inferred that the average thickness in the main part of the district may be at least 6000 ft. The extrusive rocks comprise four general types, which in ascending order of age are: andesite, rhyolite, dacite, and basalt. The veins are confined largely to the andesite, the dominant rock of the region. Characteristically this rock is augite andesite, but it includes comparatively acidic as well as basic layers, the acidic strata being

Jan 1, 1953

By Tell Ertl

THE Piceance Creek Basin in northwestern Colorado is believed to contain the richest large deposit of oil shale in North America. The major portion, about 1650 sq miles, is bounded by the White River on the north, the Grand Hogback, which runs from Meeker to Rifle, on the east, the Colorado River to the south, and the drainage basin of Salt and Douglas Creeks on the west. The Basin is a plateau 7000 to 9000 ft high, composed of 3000 ft of the Green River series of Eocene age and surrounded by a steep escarpment of 1000 to 3000 ft. From the escarpment the beds dip gently toward the center. Although oil shale is found in all members of the Green River series, the rich thick beds are found chiefly in the Parachute Creek member. This is normally capped by 100 to 200 ft of the Evacuation Creek, the youngest member of the Green River series. Actually the Piceance Creek Basin oil shale is not a shale, nor does it contain oil. It is magnesium marlstone holding solid organic matter capable of being broken down by heat into gases, some of which can be condensed into oil. The marlstone is resilient nonporous rock, difficult to sample by pick or hammer and moil. Many companies and individuals prospected the Colorado oil shale shortly after the first World War,' see Fig. 1, but because of the discovery of large quantities of petroleum in the late 1920's in Cali- fornia, Oklahoma, and East Texas, interest in oil shale development waned and did not revive until World War 11. In 1944 the writer, employed by the Bureau of Mines, prospected the high grade oil shale in numerous places along the southern outcrop of the Piceance Creek Basin. Several channel samples about 100 ft in length were cut, see Fig. 2. An eight-man crew needed a week to cut a satisfactory channel 80 ft long in the cliff face. The Phillips Petroleum Co., the Union Oil Co. of California and others also prospected the oil shale outcrop. In 1945 the Bureau of Mines initiated a core drilling program for the purpose of obtaining accurate samples from the Piceance Creek Basin. The U. S. Geological Survey set up a program for the geological study of the Basin that is still going on." Until the end of 1951 a total of 31 holes was drilled for core with diamond drills, and, in addition, cuttings through the oil shale have been obtained from 4 gas wells drilled by standard rotary oil well rigs. The Bureau of Mines drilled 12 of the holes, the Union Oil Company of California 8, the Standard Oil Company of California 6, the Sun Oil Company 3,

Jan 1, 1953

By J. P. Newell, R. W Storey

This paper describes the development of a continuous float-and-sink process to produce coal low enough in ash content to be suitable for production of electrodes. The cleaned coal had a combined iron and silicon content of 0.239 pct. The efficiency of recovery was 84 pct. AUGER mining is a form of continuous mining in that it completely replaces with a one-cycle operation the older conventional cut, drill, shoot, and load method of mining. Relatively new, having been utilized only within the last few years, it is even more continuous in its production of coal than some other forms of continuous mining being practiced in this country today. There are at present coal recovery augers in operation working along strip-mined highwalls from the outside. This paper will deal with underground mining, specifically with an experiment which was carried out with a machine developed for this purpose. The location of this experiment was in Pike County in eastern Kentucky in the upper split of the No. 3 Elkhorn seam. Coal height was 32 in. and roof conditions were very bad, with 10 in. of soft draw slate coming down with every cut in conventional mining. The coal is high-grade bituminous, excellent for metallurgical coking. Underground auger mining fills a need for a method of recovering coal in thin seams, under bad roof conditions or in split seams. In many cases it might save money by eliminating the handling of excessive quantities of rock. In fact, if the coal can be recovered, as in the case of this experiment, without mixing with any impurities outside the seam, it may be possible to load or ship a satisfactory product with little or no preparation or cleaning expense. The need for some economical method for recovering coal from thin seams is becoming more pressing as the more profitably mined seams are exhausted. Basic Principles of Auger Mining The principles of the auger, or screw conveyor, have long been used in drilling small diameter holes for blasting and other purposes. These same fundamentals apply to large diameter drilling. However, considerably more attention was given to designing a cutting head that would not pulverize the coal and consequently would require a minimum of power to drive it. The head, as designed for this job, consists of a cylinder 25 in. in diam with steel bits set in 3 positions spaced around the forward rim. A conventional coal drill bit, at the head of a bursting cone in the center of the cylinder, trails the forward rim slightly, see Fig. 1. The bursting cone and auger connections are anchored to the cutter barrel by 3 arms spaced at 120". This head design proved to be very satis-

Jan 1, 1953

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit.' The rocks are well-consolidated Gila conglomerate, quartz monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Ranaome. In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 3Y4x41/4-in. tool joints; 8-in. bits, 23/4x 3 3/4-in. tool joints; 6-in. bits, 21/4x31/4-in. tool joints; 4-in. bits, 15/ix25/s-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property. To allow for 2-in. drill pipe with tool joints, 31/2-in. core barrels and bits were used. With the standard 31h-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 5 to 37/8 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven machines. The open churn drill hole was cased with 21h-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in un-consolidated material like the formations drilled by

Jan 1, 1953

By J. L. Carne

ORE control is a matter of planning and supervision based on a foreknowledge of the content and distribution of ore. The Inspiration orebody is predominately a copper-sulphide blanket, overlain by an oxidized zone of copper silicates, copper carbonates, and barren capping. Most of the ore is treated by leaching, and the optimum requirement of this method is the goal toward which the mining schedule is aimed. Both open-pit operations and underground mining are used. The problems of ore control at Inspiration are: I—maintenance of metallurgical balance, 2—conservation of oxide ore, and 3—mining out of the orebody at an average grade which parallels that of the reserve figure. Prior to the fall of 1926 all of the Inspiration division ores were treated by flotation, but the mixed ores of the Live Oak—Keystone section carried too large a proportion of copper in the form of chrysocolla for satisfactory flotation. For this type of mixed ore the leaching process now in use had been worked out, and the plant was put into operation in the fall of 1926. The process is known as the ferric sulphate leach, and the leaching solvents contain both ferric sulphate and free sulphuric acid. Sulphide copper in the form of chalcocite is leached by the ferric sulphate, and the oxidized copper minerals are dissolved by the sulphuric acid. From the beginning of operations to the present time the plant has treated some 67 thousand tons of ore having an average copper content of 1.142 pct, of which 0.605 pct has been in the form of oxidized copper minerals. The plant is currently being operated at the rate of 11,500 tons per day, 7 days per week. However, this tonnage is produced in 6 days of mine operation. The average current grade is approximately 1 pct copper, of which about 50 pct is present as oxidized or acid-soluble copper. Close control of metallurgical procedures is a vital factor in operation of the plant and is in turn dependent upon careful regulation of feed. An excess of sulphide copper requires the presence of more ferric sulphate in the leaching solvents. On the other hand, an excess of ferric sulphate in the presence of a low sulphide feed seriously reduces the efficiency of electrolytic precipitation in tank house operations. This condition must be avoided at all costs. An additional factor must be considered. It is known that the sulphide content is predominant in the ore which remains in the reserve. To maintain the necessary sulphide-oxide proportion, so that the entire reserve will mine out at both a grade and a sulphide-oxide ratio that will be permissible throughout the remaining life of the property, it is obvious that the oxide reserve must be currently preserved to as great an extent as possible. Fortunately much of the oxide reserve is found in the open-pit operation. This happy circumstance adds flexibility to the procedure and makes it possible to maintain the close regulation necessary to insure best overall results, both mining and metallurgical. The two ore streams from the pit and underground operations join at the coarse-crushing plant, and proper control must be exercised before the ore reaches this point. The open pit is currently providing about 55 pct of the ore and underground operations the balance of production. Both operations are under a single superintendent so that the closely interrelated problems may be more readily controlled. The least flexible of the two producing units is the underground mine, where the mining method is block caving. Although a block of ground may be undercut and developed, characteristics of the material determine its caving action. The ore in the chutes must be drawn. The area currently mined from underground is in a zone predominately sulphide and is generally

Jan 1, 1953

By F. R. Zacher

Unmined blocks in the Sewickley seam, surrounded by worked out areas, have been found to transmit overburden oads through the interval strata to the Pittsburgh seam workings 90 ft below. Operating experiences under these conditions are described in this paper. IT has always been understood, in northern West Virginia where both the Pittsburgh and Sewickley seams are mined, that pillaring or splitting in the lower Pittsburgh seam could break the interval strata and make mining in the overlying Sewickley difficult or impossible. In some instances, however, mining methods practiced in the overlying Sewickley have reversed the problem, resulting in a shifting of the intermediate layer and destroying valuable acreages of the Pittsburgh seam below. The Pittsburgh seam in this area is, on the average, 102 in. thick. About 18 in. of low-quality head coal are left for roof protection as it is mined. Immediately over the coal are two draw slates, two rider seams, and a good hard black shale, locally called black rock, which serves as a permanent roof where draw slates and riders are taken down on certain main haulroads. In most areas mined in this seam in the past the bottom has been satisfactory but soft. However, in areas now being mined by the Christopher Coal Co. in Monongalia County the bottom is very hard and unyielding. The Sewickley seam, lying 85 ft above the Pittsburgh, will average nearly 70 in. in thickness and is overlain by massive shales and sandstones. Much of the seam is mined only down to the "sheep skin," a hard band about 4 in. above the bottom, which is composed of relatively hard shale. Fig. 1 shows a typical cross-section through the Pittsburgh and Sewickley seams in the area being discussed. The thin Redstone seam is not deep-mined but is stripped in some locations where it outcrops. Most of the Sewickley seam in the Morgantown areas has been mined out from 10 to 30 years. There are at present only two large mines and five or six others of small production operating in that seam. Mining systems in the past were generally of the room and pillar type with some full pillar extraction. This seam in the Scott's Run area was operated by as many as thirty or forty different companies at one time, during an era beset with strikes, market failures, and depressions. Most of the barrier pillars along the many main haulroads were abandoned, and blocks of Sewickley coal that would be fully recovered with today's machinery were left unmined when conditions became anything less than favorable. Consequently the mined-out Sewickley is spotted throughout with solid blocks varying in size from 100x100 ft to 200x400 ft and larger. It is these unmined Sewickley blocks surrounded by worked-out portions that have caused transmissions of roof stresses to underlying Pittsburgh mining areas, see Fig. 2, resulting in untold difficulty. As the Christopher Coal Co. operates both the Pittsburgh and Sewickley seams, accurate maps of workings in both seams are readily available and are regularly consulted. The first Pittsburgh squeeze attributed to overlying Sewickley blocks occurred in the Arkwright mine, where a shuttle car crew was developing a section consisting of seven panel headings. These entries were being driven in solid coal 3400 ft from the nearest worked-out Pittsburgh area. As mining advanced it was noticed that certain ribs began to slough and certain areas to show signs

Jan 1, 1953

By P. A. Laxen, H. R. Spedden

WHEN a mineral particle is fractured, bonds between the atoms are broken. The unsatisfied forces that appear at the newly formed surface are considered to be responsible for the adsorption of ions at the mineral surface. A knowledge of the mechanism and extent of ion sorption from solution onto a mineral surface is of interest in the development of the theory of flotation.'*' Study of the adsorption of sodium from an aqueous solution oftheon quartz offers a simple approach to this complicated problem. The availability of a radioisotope as a tracer element meant that accurate data could be obtained."." Three main factors which appeared likely to affect the adsorption of sodium are: l—concentration of sodium in the solution, 2—concentration of onotherof cations in the solution, and 3—anions present in the solution. Hydrogen and hydroxyl ions are always present in an aqueous solution. By controlling the pH, the concentration of these two ions was kept constant. The variation in thesethe amount of sodium adsorbed with variation in sodium concentration was then determined under conditions standardized in regard to hydrogen ion. The effect of concentration of hydrogen ions and of other cations was also measured. A few experiments were made to get a preliminary idea on the effect of anions. The active isotope of sodium was available as sodium nitrate. Standard sodium nitrate solutions were used throughout these experiments except when the effects of other anions were studied. It was found that sodium adsorption increased with sodium-ion concentration, but less rapidly than in proportion to it. Increasing hydrogen-ion concentration, or conversely decreasing hydroxyl-ion, brings about a comparatively slight decrease in sodium-ion adsorption. Increasing the concentration of cations other than hydrogen or sodium decreases somewhat the adsorption of sodium ion. It would appear as if the kind of anion is a secondary factor in guiding the amount of sodium ion that is adsorbed. Materials and Methods Quartz The quartz was prepared as in previous work in the Robert H. Richards Mineral Engineering Laboratory' except for the refinement of using de-ionized distilled water for the final washing of the sized quartz, prior to drying." To minimize the laborious preparation of quartz, experiments were made to determine .whether the sodium-covered quartz could be washed free of sodium and re-used. The experiments were successful as indicated by lack of Na" activity on the repurified material and by its characteristic sodium adsorption. Table I gives the spectrographic analyses of the quartz used. The quartz ranged from 16 to 40 microns in size, averaging about 23 microns (microscope measurement), and had a surface of 1850 sq cm per g (lot I), 2210 (lot 11) and 2000 (lot 111) as determined by the Bloecher method." Radioactive Sodium Method of Beta Counting for Adsorbed Sodium: Na22, the radioisotope of sodium, possesses convenient properties.' It has a half-life of 3 years, thus requiring no allowance for decay during an experiment. On decay it emits a 0.575 mev ß+ radiation and a 1.30 mev r radiation. The decay scheme is illustrated in the following equation: ß+ NaR-------'8'77NeZ2 3 years The /3 radiation is sufficiently strong to penetrate an end-window type of Geiger-Mueller counting tube. This, in turn, makes it possible to use external counting, a great advantage in technique. Furthermore, it permits the assaying of solids arranged in infinite thickness, while assaying evaporated liquors on standardized planchets. The equipment used was standard and similar to that employed by Chang.R The original active material was 1 ml of solution containing 1 millicurie of Na" as nitrate. This active solution was diluted to 1000 ml. Five milliliters of this diluted active solution was found to give a quartz sample a sufficiently high activity for accurate evaluation of the sodium partition in the adsorption measurements. Also, 1 ml of final solution gave a sufficiently high count for precision on the liquor analyses. The sodium concentration of the diluted active solution was 1.2 mg per liter, so that 6 mg of sodium for 60 ml of test solution and 12 g of quartz was the minimum amount used. The active solution was stored in a Saftepak bottle. Procedure for Adsorption Tests: The method consisted of agitating 12 g of quartz with 60 ml of solution of known sodium concentration for enough time to establish equilibrium between the solution and the quartz surface. The quartz was separated as completely as possible from the solution by filtering and centrifuging. The activity on the quartz and in the equilibrium solution was measured and the partition of the sodium was calculated from the resulting data. The detailed procedure for the adsorption test is set forth in a thesis by Laxen." In brief, it included the following steps: 1—Ascertainment of linearity between concentration of Na" and activity measured. 2—Evaluation of factor to translate activity on solid of infinite thickness in terms of activity on an evaporated active film of minute thickness, on the various shelves of the counter shield. 3—Taking precautions to avoid evaporation of water during centrifuging

Jan 1, 1953

By Norman F. Schulz, Emert W. Lindroos, Strathmore R. B. Cooke

DURING the course of an investigation of the effects of various starch products on hematite and quartz in regard to their separation by flotation, it was found that whereas most starches flocculated suspensions of hematite in water, they did not flocculate similar suspensions of quartz. However, a derivative of whole corn starch containing approximately one tenth aminoethyl group (-—CH,CH2NH2) per glucose unit proved to be an exception. This aminoethyl starch was a very good flocculant for both hematite and quartz, and also possessed the property of depressing both minerals in cationic flotation tests. In view of the widespread use of starch as a flocculating agent of solids in tailing water, pulps, and for coal slimes, ore and various other purposes in sulphide and nonsulphide flotation, attention is called to the effective surface activity of aminoethyl starch in flocculating suspensions of the two minerals investigated. Test Materials The starch products employed included Globe Pearl starch No. 144 and fresh aminoethyl starch, both supplied by the Corn Products Refining Co. Dr. A. L. Elder" has kindly furnished a description The minerals used in the various tests are listed in Table I. Their specific surfaces were determined by the air permeability method, which gave results reproducible to within plus or minus 3 pct. Surface areas obtained by krypton adsorption were greater than the air permeability values by a factor of 2.0 for quartz and 62.4 for hematite. Starch Adsorption on Mineral Surfaces Starch adsorption data were obtained by agitating for 1 hr measured volumes of neutral starch solutions of known concentrations with weighed quantities of finely divided mineral of measured specific surface. The mineral then was separated from the pulp liquor by centrifugation and the residual starch content determined by chemical analysis. The analytical procedure consisted of oxidizing the starch with an excess of dichromate in acid medium at elevated temperature and titrating the excess dichromate by the ferrous ammonium sulphate—potassium permanganate method. Analytical results were reproducible to within less than 1 pct. Adsorption data obtained in unbuffered solutions near the neutral point are plotted in Fig. 1. The adsorption on either mineral was far greater for the aminoethyl starch than for the Pearl starch. Flocculation The degree of flocculation was determined for a given pulp by measuring the time required for the supernatant liquid to reach a certain degree of clarity. This was accomplished by passing a standard light beam through the pulp, contained in a glass settling tube, and measuring the time required after agitation ceased for the intensity of the transmitted beam to reach a predetermined value. Starch reagent was added to each pulp in small uniform increments, and the settling time was determined after each such addition. Increase in pulp liquor volume due to the added starch reagent was ignored in these tests. Pulps were maintained at or near neutrality. of the preparation of the aminoethyl starch: "Three hundred and sixty grams of powdered unmodified corn starch was slurried in 3 liters of water and stirred during gelatinization. Eighty-six grams of ethylenimine was added slowly during reflux which was continued for 4 hr. The yellow solution set to a gel upon cooling. This was broken up and neutralized with HC1. The thick paste was dried on hot rolls. Adsorbed polyethylenimine and salts were removed from the yellowish products by washing first with acidified aqueous '60 per cent' alcohol. The white product contained nitrogen corresponding to about one aminoethoxy group per ten glucose residues." Only laboratory samples of this starch derivative have been prepared. The starch reagents for use in flocculation, flotation, and adsorption tests were made up by dispersing the appropriate quantities of air-dry starch in cold water, adding boiling water, and heating for 30 min at 120°C under pressure. When used ip adsorption tests, the starch solutions first were cleared of insoluble matter by centrifugation.

Jan 1, 1953

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit.' The rocks are well-consolidated Gila conglomerate, quartz monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Rannome. In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 31/4x41/4-in. tool joints; 8-in. bits, 23/4x 33/4-in. tool joints; 6-in. bits, 2Y4x3Y4-in. tool joints; 4-in. bits, 15/ix25/8-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property. To allow for 23/8-in. drill pipe with tool joints, 31h-in. core barrels and bits were used. With the standard 31h-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 55/8 to 3 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven machines. The open churn drill hole was cased with 21h-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in un-consolidated material like the formations drilled by

Jan 1, 1953

Jan 1, 1953