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By O. J. Huber

HYDROGEN effects in titanium alloys have been the subject of extensive research in recent years. Lenning, Craighead, and Jaffee1 showed that hydrogen embrittles a titanium and, at the same time, elevates the temperature at which the transition from brittle to ductile failure occurs during impact testing. In later work: the same investigators concluded that several all-ß alloys tolerate large amounts of hydrogen without serious damage to mechanical properties, but two-phase alloys suffer varying degrees of embrittlement, depending upon the specific ß-stabilizing addition used and whether or not an a stabilizer is also present. The mechanism of hydrogen embrittlement has not been clearly established. Strain rate has been shown to be an important factor and, therefore, a strain-aging phenomenon is suspected to cause the effects that hydrogen has produced on mechanical properties.3-5 Ripling6 has pointed out that ductility damage varies inversely with strain rate in a-ß alloys and directly in a alloys. This effect in a alloys can be associated with transition-temperature behavior, but the embrittlemeat mechanism in two-phase alloys remains in question. During the course of research conducted for the U. S. Air Force, a dark-etching phase has been observed at the grain boundaries of certain two-phase titanium alloys after duplex (solution plus aging) heat treatments.5-7 An example of the dark-etching interface phase is shown in Fig. 1. Decrease in ductility accompanies the appearance of the minor phase but has not always been attributable solely to the presence of the phase. Although the dark-etching phase has been observed in alloys under a rather narrow range of heat-treatment conditions, it is of general interest since it is always associated with the presence of hydrogen. The known embrittling effect of hydrogen makes any facet of its behavior of particular importance. The intergranular phase has been observed in titanium-base alloys with relatively high amounts of the ß-stabilizing elements. Specific examples of such alloys include Ti-5 pct Mn-2.5 pct Cr, Ti-3.5 pct Cr-3.5 pct V, Ti-8 pct Mn, and Ti-3 pct Mn-1

Jan 1, 1958

By H Chen

"An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation. The Mesozoic iron oxide-copper-gold (IOCG) mineralisation in southern Perú and the northern Chile coastal IOCG belt have emerged as two of the major exploration targets in the Central Andes in the last two decades. These Mesozoic Andean IOCG deposits formed in two mineralisation epochs, the Middle-Late Jurassic (170–150 Ma) and Early Cretaceous (130–110 Ma), with the major Cu-rich IOCG deposits located in the Early Cretaceous mineralisation belt. The early studies in some IOCG centres support a magmtic-hydrothermal model for both Cu-rich IOCG deposits and broadly contemporaneous Cu-poor ‘Kiruna-type’ magnetite deposits. However, recent evidence from field, geochronology, fluid inclusions and stable isotopes show that in the major Central Andean deposits (eg Raúl-Condestable, Mina Justa in southern Perú, La Candelaria-Punta del Cobre and Mantoverde in northern Chile), the incursion of evaporite-sourced basinal brines or sea water may be a prerequisite for economic Cu mineralisation. Moreover, the Cu-poor massive magnetite deposits could be interpreted as a product from iron oxide melt, with minimal intervention by non-magmatic fluids. Therefore, they could represent two contrasted deposit types and contain little genetic relationships. The different roles of magmatism during the Mesozoic are responsible for the variable mineralisation in the Central Andes, including IOCG, magnetite and ‘manto-type’ deposits. The ore modelling for both the Middle-Late Jurassic and Early Cretaceous metallogenic epochs in the Central Andes are reconstructed.CITATION:Chen, H, 2015. The Mesozoic iron oxide-copper-gold (IOCG) mineralisation in the Central Andes – a refined IOCG ore-forming model in the Palaeozoic Continental Margin, in Proceedings PACRIM 2015 Congress, pp 251–254 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 18, 2015

By Bernard Baum

Combined preservative/fire retardant systems were pressure impregnated into Douglas fir and Ponderosa pine samples and evaluated for decay, flammability, and mechanical properties. The objective of the evaluation was to identify cost-effective systems for use with mine timber, and to provide an information and data base from which Federal code regulations could be established. Thirty systems, based on combinations of commercially available preservatives and fire retardant impregnants and coatings, were applied to Douglas fir samples and evaluated for decay and flammability. The ten most promising systems were ap¬plied to both fir and pine and reevaluated for decay and flammability before and after water leaching. The five best systems were given a thorough laboratory evaluation using both pine and fir. Evaluations included flame spread, smoke density, toxic gases, decay, compressive strength, flammability after artificial weathering, and cost. All systems had acceptable toxic gas levels and smoke densities, but flame spread was only borderline acceptable for most of the samples tested. A trade-off analysis showed Acid Copper Chromate/Fyrprufe with Albi-107A/144 intumescent paint and Koppers NCX to be the-best systems. Recommendations are made for further large-scale testing using the two best Systems.

Jan 1, 1978

By D. Cronin, A. Plumtree, A. Hosseini

"Crack in corrosion (CIC) linepipe defects are hybrid defects not directly addressed by existing assessment methods. To investigate the collapse pressure of these defects, CIC defects were formed in 508 mm (20 in) diameter, 5.7 mm thick API X52 linepipe. These defects were 200mm long and varied in depth from 3.0 to 3.8 mm. The respective failure pressures decreased from 14.3 MPa for the plain pipe to 7.9 and 6.2 MPa respectively. A standard FAD method was applied to evaluate these defects as equivalent length and depth cracks, predicting the rupture pressure conservatively by 18% on average. However, when assessed as equivalent cracks with a reduced wall thickness corresponding to the corrosion depth, the predicted collapse pressures were more accurate, approximately 14% conservative on average.INTRODUCTIONPipelines are a safe and efficient method for transporting high volumes of oil and gas. However, aging pipelines may experience defects, such as cracks or corrosion. Present approaches include detection, assessment and repair of defects before they may become critical to the integrity of the line. There are different codes and standards for assessing pipeline defects, depending on the type of defect, with API 579 (API, 2000) being the most common code for the assessment of cracks in oil and gas pipelines . Corrosion is commonly assessed using RSTRENG (Kiefner & Vieth, 1989). Other methods such as the Finite Element method and the CPS method (Cronin & Pick, 2002) may be used for crack and corrosion, respectively.At present, there is no specific method or procedure for assessing crack in corrosion (CIC) defects, and as a consequence there is a need to provide an assessment and evaluation of the integrity of these defects in order to determine the requirements for defect repair or hydro line testing. Hence, the goal of this study was to provide guidance for the assessment of CIC defects for linepipe.CRACK IN CORROSION DEFECTS (CIC)Significant amounts of corrosion and cracking are not generally found together (Hosseini, 2010). However, a new hybrid defect including cracking within corrosion (CIC) has been identified in operating pipelines. In this case, cracking coincident with corrosion (schematic shown in Figure 1) provides the motivation for the current study."

Jan 1, 2012

"Dry crushing trials at pilot scale have shown that the IMPTEC Super-fine Crusher has the capacity to generate ultra-fine products from hard mineral assemblages presenting at either coarse or fine feed sizes. The super-fine crusher was introduced at the 27th International Mineral Processing Congress in Santiago Chile, with results from a range of duties, which demonstrated the crusher’s potentially expansive operating scope. This paper focuses on current preparations to progress the technology from pilot scale to commercial scale within a concrete production plant, where lead blast furnace slag will be crushed to ultrafine sizes to generate an active silica product for use as a supplementary cementitious material, which can replace Ordinary Portland Cement. A key objective of the operation is to demonstrate an economical route to reducing the cost of concrete production, improving concrete quality and reducing greenhouse gas emissions. The paper also highlights the potentially key contribution super-fine crushing can make to the long-term viability of iron ore beneficiation by simplifying comminution flow-sheets and reducing operating costs. There is every indication that the technology has the potential to feature universally in metalliferous and industrial mineral flow-sheets. INTRODUCTIONToday’s IMPTEC super-fine crusher is the successful end product of development work over 19 years and is the third particle size reduction device conceived, built and tested over that time. The concept evolved as engineering design progressively developed a mechanism capable of harnessing a highly efficient combination of fundamental particle breakage mechanisms. Early development trials at pilot scale indicated that super-fine crushing had the potential to establish a wet or dry, media free, low energy route to ultra- fine products from hard mineral assemblages. Having achieved some impressive results from open circuit pilot tests on fly ash and granulated blast furnace slag it became clear that the cement industry may well provide opportunities to apply the technology commercially. Following expressions of interest from local cement and concrete plant operators, IMPTEC negotiated an agreement to undertake commercial trials with a SFC 350 super-fine crusher installed on site at a concrete supply plant in South Australia. Super-fine crushing scope was recently expanded following further expressions of interest from research groups involved in the processing of iron ores. Pilot trials are progressing in these areas."

Jan 1, 2016

By J. F. Myers

THE Tennessee Copper Company's operations are in the Ducktown Basin, in the extreme southeast corner of Tennessee. The ore is of the heavy sulphide type, the predominating sulphides being pyrite, pyr-rhotite, chalcopyrite and sphalerite. The gangue is a micaceous schist, with varying amounts of free quartz. The mining operation develops two classes of ore, each of which is concentrated by the flotation process in separate mills. Table 1 shows a typical analysis of the ore treated in each mill. TABLE 1.-Typical Analysis of Ore Treated Composition, Per Cent Mill Tons Capacity Cu Fe S Zn London 1,350 1.41 28.4 20.2 1.10 Isabella 850 0.69 44.7 31.6 0.31 LONDON MILL The London mill was built in 1922, with a capacity of 400 tons per day. This was gradually increased to 600 tons by 1926, to 900 tons in 1928, and to 1350 tons in 1938. Fig. 1 shows the crushing and grinding flowsheet in 1928, when the capacity was 900 tons; Fig. 2, the crushing and grinding flowsheet that was developed by 1938, for the treatment of 1350 tons per day. A comparison of these flowsheets reveals that this increase in tonnage, with a subsequent improved operation, was accom-plished with identically the same major equipment. Improving a concentrator's capacity, metallurgical results, and costs as time goes by is a commonplace practice, and is accomplished in various ways, depending upon the local conditions. The reason for presenting the data of how this was accomplished in the mills of the Tennessee Copper Co. is that a fast-running rod mill plays such an interesting role

Jan 1, 1940

By Fengming Li, Baocheng Song, Yuemin Zhao, Hailin Dong, Liping Peng, Chusheng Liu, Chenlong Duan

The large-scale vibrating screens with high processing capacity and efficiency are now in great demand in the mineral processing industry of China. However, as the equipment is always suffering from the strong external harmonic exciting force, the inertial force and the impact from the minerals, the fracture of the beam and the crack of the lateral plate are easy to occur, which has a negative effect on the reliability of the equipment itself and the industrial production. Considering the essential high reliability of the equipment, a new large-scale vibrating screen with duplex statically indeterminate mesh beam structure (VSDSIMBS) was developed. The natural characteristics, namely, the natural frequency (or the modal frequencies) and the natural mode, and the dynamic characteristics of the VSDSIMBS were obtained by the modal analysis and the harmonic response analysis. In addition, the dimensions and the constructions of the statically indeterminate plate and the mesh beam were chosen as the design parameters, to which the sensitivity of the whole mass, the maximum deformation and the maximum stress of the duplex statically indeterminate mesh beam structure (DSIMBS) were summarized for the optimization of the DSIMBS. Now the large-scale VSDSIMBS is being well used in China.

Jan 1, 2014

By Vinay Kumar

The disposal of large quantities of electronic scraps generated world wide is causing not only environmental problem but also loss of resources. Therefore, R & D efforts have been made to develop a suitable process for extraction and separation of metals viz. copper, zinc, cadmium and nickel expected from the leaching of e-scraps using solvent extraction process. Different process parameters viz. pH, aqueous to organic ratio, kinetics of extraction, extraction isotherm to establish stage requirement for extraction in continuous mode etc have been studied to optimise the condition for the separation of metals. The studies showed the selective extraction of copper after iron precipitation from the sulfuric acid leach solution of e- wastes containing ~1.0 g/L Cu and minor impurities with 2-hydroxy-5- nonylacetophenoneoxime (LIX 84) above pH 2.0 in single stage. A scheme for separation of cadmium, zinc and nickel has been proposed using di(2-ethylhexyl) phosphoric acid (D2EHPA) under controlled pH of the aqueous solution. The order of extraction was found to be Zn> Cd> Ni.

Jan 1, 2009

By Walter E. Lewis

A mothod of cutting St. Poter sandstono by streams of water under pressuro from nozzles and pumping the resulting, sand to the surface has roduced the cost of storm-sowor construction in the Minneapolis-St. Paul aroa. The hydraulic tunneling method, has boon in uso since 1934, and it is adaptable only to the lower part of tho St. Potor, sandstone, which lios within or just abovo the ground-wator levol. Tho upper part of the formation, immediately bolow tho Plattoville limestone, 18 comonted and camot bo mined by hydraulic mothods. The tunnel oporation observed in this instance is a main storm sowor starting at tho Mississippi Rivor below the 10th Avenuo SE bridge and will be drivon 6,000 foot at a grado of 1 foot por thousand undor the city of Minneapolis to Konnody and Arthur Stroots NE. A lateral storm sowor, which branchos from the main sovor at 8th Street SE, will oxtond from 10th Avonuo SE to 15th Avonuc SE. Tho tunnele are being drivon with the floor approximately at ground-wator lovol. All waters ontoring the main storm sower vill drain into the Mississippi Rivor. Tho main storm sewer cross soction is excavated to a roughly hox- agonal shapo 10 foot high by 10 foot wido at the contor. The upper half of the tunnol soction is solidly lined with either timber or stool. After all oxcavation is completed, the tunnel will be finished with a concrete circular lining 8 feet in inside diameter. The lateral storm sewer is excavated to a 5-1/2-foot by 7-1/2-foot section and will event- ually be finished with concrete to a modified horseshoe-shaped section 6 feet by 4 feet. Where the back or walls are considered unsafe during excavation, the upper half of the tunnel is lined solidly with timber. The equipment used by the Sower Division, City of Minneapolis, has been selected to meet a number of applications in storm-sewer work with diversified usage in mind. The sand pumps used on this job are rubber-lined, designed to work against a head of 100 foet. Either 4-inch or 6-inch diameter suction and discharge openings may be used. Other equipment, such as pipe, mine rail, flexible hoses, compressors, and general surface shop equipment, is drawn from a contral warohouse. It may be either new or used, depending on what is available at the warehouse. Much of the pipe and mine rail has been used for many years on several different tunnol jobs.

Jul 1, 1950

By Robert P. Koenig

FOR the first time other than on occasion of war the people of the United States are experiencing full-scale participation in world affairs. Public concern has seldom been so involved with conditions and events beyond our own continental limits. Our interest is not wholly of recent origin. Over the years the activity of the United States as a world power and an important trading nation brought about the widening of our political interests and the progressive interrelation of our economy with that of the rest of the world. Some may see in our present phase of world-wide activity the natural consequence of our continued growth. Such observers must also be aware that within recent years our importance in world affairs has been considerably increased. For within the space of an extremely short period events have occurred that have literally jet-propelled the United States into the role of a leading participant in the conduct of world affairs. Little of the choice of this role was left to us. The decline of power in western Europe, the ascendancy between the Elbe and the Pacific shore of an obviously antagonistic force, our earnest endeavor to secure international control of atomic energy, the necessity of providing relief and rehabilitation to war-ravaged areas, and the police work that must be done in occupied territories are problems which have risen in great measure without our help-but they cannot be solved without it.

Jan 1, 1948

By D. J. Pye

Certain high molecular weight synthetic polymers in very dilute solutions decrease water mobility in porous media 5 to 20 times more than would be expected from the solution viscosity. This indicates that increasing water viscosity to reduce adverse mobility ratios in water floods is economically feasible in many situations. Laboratory results are consistent with accepted theory with respect to improved areal sweep, permeability distribution and displacement efficiencies as would be predicted from the solution mobility measured in the core. The laboratory results have been confirmed by field pilot floods which recovered 80 per cent and 100 per cent more oil than comparative water-flood pilots in fields containing 16 cp and 130 cp oil respectively. Economic analyses based on pilot data indicate the method to be profitable. INTRODUCTION The desirability of improving the water-oil mobility ratio in waterflood-ing operations has long been appreciated, especially for high viscosity crudes. Methods of reducing the oil viscosity by application of heat, gas resaturation and miscible drives have met with some success. Conversely, it has been generally conceded that increasing the water viscosity would also be effective, but the amount of thickening agent required to effect an appreciable increase in viscosity is discouragingly high. The use of such materials as glycerin, sugar, or glycols is completely out of the question economically, so that considerable attention has been given to the use of the much more efficient synthetic water-soluble polymers as viscosity improvers; but even these materials are not economically attractive. If increased water viscosity could be economically realized, marked improvements in areal sweep efficiency, as discussed by Caudle and Wittel and others, would be realized in field operations. Similarly, the improved mobility ratio would bring about increased recoveries by correcting permeability distribution problems as discussed by Stiles2 and Dykstra and Parsons,' and improvements in displacement efficiency would be obtained as recognized by Buckley and Leverett4 and Welge.5 LABORATORY STUDIES THE RESISTANCE EFFECT In the course of research to develop a more efficient water-soluble polymer viscosity builder, we discovered that a very few types of water-soluble polymers exhibited a most unusual and very interesting property not previously observed. Among the earliest materials showing such activity were polymers containing acry-lamide. Fig. 1 shows a typical plot of the solution viscosity vs concentration for one of these polymer solutions as determined in an Ostwald viscometer. As usual, such a plot is an approximate straight line on semi-logarithmic paper. Now, an inspection of Darcy's law, shows that a viscosity value can also be determined in a formation sample if the pressure, flow rate and permeability are known. If the core permeability is previously determined with wa- ter or brine, then the viscosity of another aqueous fluid may be determined from the equation. The viscosities so determined for glycerin (glycol or polyvinyl alcohol solutions, for example) agree with the viscometer determinations. However, in the case of the water-soluble polymer solutions in which we are interested, the viscosities measured in the formation sample depart very markedly from the viscometer values. A typical relationship for a polymer solution is also shown in Fig. 1. This unusual departure from the expected response is referred to as a resistance property of the polymer and is quantified as "resistance factor". Since the resistance factor R can best be determined in a core under study, the term has been defined on the basis of the ratio of the brine mobility to the polymer solution mobility at residual oil saturation so that:

Jan 1, 1965

By T Baumgartl, R Akesseh, M Edraki, B Usher, M Shaygan

With the continuing demand for industrialisation and urbanisation, the rate of mining and processing lower grade ores will also increase, producing large quantities of waste. As extreme weather events associated with global climatic changes are to be expected, the risk from mine tailings management will also increase (Edraki et al., 2014). In sulfide rich tailings, weathering within the critical zone controls the fate and distribution of the contaminants such as copper, zinc, and lead in the environment (Hayes et al., 2014). Tailings evolve geochemically and physically with time as they age under the influence of complex interacting factors. Understanding of the complex interactions between these controlling factors is important for predicting and mitigating the potential risks associated with sulfide tailings such as generation of acid mine drainage (Lindsay et al., 2015). Wetting and drying initiates weathering processes, influencing the hydrological, geochemical and geomechanical properties of the tailings. In semi-arid climates where evaporation exceeds precipitation, salts may precipitate within the pore system as well as on the tailings surface. These salts become the temporal sink of heavy metals until the next episodic rainfall (Nordstrom, 2011, Grover et al., 2016). Therefore, using kinetic and static test to describe the trajectory of these material evolutions without considering the complex interactions could potentially lead to erroneous prediction. In order to understand these relations, the effect of drying and wetting cycles on the hydrological, geochemical and geomechanical properties of ageing tailings was studied to draw conclusions for advanced tailings site rehabilitation and mine closureCITATION:Akesseh, R, Shaygan, M, Usher, B, Edraki, M and Baumgartl, T, 2018. Wetting and drying cycles as a means to determine the effect and consequences of tailings ageing, in Proceedings Life-of-Mine 2018, pp 24–27 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 25, 2018

Transport Canada (Federal Government) constructed the Halifax International Airport (HIA) near Halifax, Nova Scotia, Canada in the late-1950s. Site location criteria were primarily; direct access to the city and a raised flat topography. Unfortunately, the site chosen was located on shallow Meguma group pyritic slate bedrock. In 1982, during the construction of an airport taxiway, it became very apparent that acidic run-off from excavated bedrock at the site was having a negative impact on adjacent aquatic habitat. The ensuing 20 years involved many efforts to rectify the acidic drainage problem with mixed results. Some of the techniques which have been implemented at HIA to address the acid rock drainage (ARD) issue include: dry covers, wet covers, wetlands, electrochemical treatment, open limestone channels, limestone gravel polishing beds, agricultural limestone application, and two hydrated lime [Ca(OH)2] addition treatment plants, the first producing low density sludge and the new one that creates high-density sludge. Based on over 20 years of water quality monitoring data it is apparent that only certain acid drainage mitigation methods have been effective at HIA. In 2000 when the Halifax International Airport Authority (the Authority) took control of the operation and management of the airport, including the ARD issue, it was decided to construct a modern, highdensity sludge, hydrated lime treatment system. This new $CAN 7 million system is now in operation and is achieving compliance with water quality guidelines. Due to high operating cost, a review of methods to optimise the system is being carried out. The HIA ARD problem, though minor by comparison to most mine sites, has provided unique challenges and may provide valuable insights into this urgent environmental and financial issue.

Jan 1, 2003

By D. J. Zhao

The reduction and nitridation of preoxidized ilmenite pellets were investigated with the prospect of creating grain refiners containing titanium oxides and titanium nitride dispersoids. All the reduction and nitridation reactions were conducted in a furnace with a thermogravimetric analysis system under a H2-N2 gas mixture of 5 L/min. The ilmenite pellets were reduced at 1100°C for 4 hours under 80%H2-20%N2 gas mixture, and then respectively nitrided at 700°C, 800°C, 900°C and 1000°C for 12 hours under 10%H2-90%N2 gas mixture. The products were characterized by X-ray diffraction (XRD) and Scanning Electron Microscopy (SEM), and analyzed by Energy Dispersive Spectrometers (EDS) and Electron Probe Mapping Analysis (EPMA). Iron oxides were reduced into metallic iron, and oxides of titanium were found, but titanium nitride particles were not identified in the products under the prevailing conditions.

Jan 1, 2014

By G. W. Crane

When ore was discovered on the Tintic Standard property in the spring of 1916, new developments were immediately started both north and south of that property, on the supposition that in East Tintic the ore would be found to occur in northerly trending channels, similar to those so characteristic of the older camp. But 9 years of developments have failed to establish the existence of such a channel. This is due undoubtedly to important differences in the geology of the two camps, that were not recognized, and is indicative of the danger in the general tendency toward interpreting the geology of East Tintic too strictly in terms of that of the older camp. It should not be forgotten that while the two sections have had practically the same geological history, the net results of a long series of geologic processes have developed in them conditions that differ so widely and in so many particulars that the necessity of regarding them as essentially different districts seems apparent. In an effort to emphasize this fact, this paper is devoted chiefly to a comparison of the geological conditions as found in the two camps and is addressed more particularly to those familiar with the geology of the Tintic district. Following a short description of the several ore discoveries and of the Tintic Standard mine is a brief discussion of the evidence of ore channels in East Tintic. For my information regarding the Tintic Standard mine, I am indebted to the courtesy and generous assistance of Messrs. Wade and Snow of the Tintic Standard Mining Co. East Tintic District The Term East Tintic district is here used to represent that eastern part of the Tintic mining district lying between the productive portion of the old camp of Tintic on the west, the foot hills of Goshen Slope, on the east, the Denver and Rio Grande Railroad on the north, and Silver Pass on the south, all in Utah County. As represented by the geological map, Fig. 1, it includes an area about 4 miles wide, east and west, and 5 miles long, north and south, within which the Tintic Standard mine is about centrally located. The map is a compilation of data supplied, for the most part, by the Chief Consolidated Mining Co. and represents the

Jan 1, 1927

By James B. Pfeiffer

TETRA Technologies, Inc., has applied its HDS? metals precipitation system to control a mine drainage problem. The effluent from the mine contains primarily zinc, manganese and iron, plus varying small amounts of cadmium, copper and lead. The flow from the mine averages 350 gpm, and varies considerably depending on season and weather conditions. The plant was designed to treat 1000 gpm of the water and generate a dense sludge suitable for recycling. The HDS? system contains a number of reactors, followed by post filtration and meets stringent metals discharge criteria as specified by New Source Performance Standards.

Jan 1, 1994

By P. R. Smith, R. A. Walters

Borates have been mined in the desert areas of California and Nevada for more than 100 years. To about 1890, playa surface mining provided the chief sources of boron minerals. Underground mining of colemanite and later of borax and kernite was predominate until about twenty years ago. Open pit mining of the large deposits of borax and kernite near Boron, California has been most significant for the past twenty years. Mining of colemanite in the Ryan, California area, near Death Valley, began in 1907. Following the discovery of the large deposits in the Boron area (about 1957), mining in the Death Valley area became nearly nonexistent. Only small tonnages were mined for special uses. Little mining was done in the Boraxo area near Ryan. The first claim was made in about 1915. In 1960 the area became the property of the Kern County Land Company, which was acquired by Tenneco Oil Company in 1967. In 1976 the various boiate properties and claims in this region were acquired by American Borate Corporation. The open pit mine is now approximately 122 m (400 ft) deep, 910 m (3000 ft) long and 305 m (1000 ft wide). The borates in the Boraxo pit consist primarily of three minerals. These are about 50% colemanite (CB6011 5H20), about 40% probertite (NaCaB50g 5H20), and 10% ulexite (NaCaBgOg 5H20). The colemanite, along with boric acid and high-grade colemanite ore from Turkey provide the only sodium-free borates for production of textile grade fiberglas. When heated to its decomposition temperature, colemanite decrepitates to a fine powder, which is the basis for the concentration process. The gangue minerals in this deposit are primarily calcite and clays, including bentonite. The ore body has a very low arsenic content, which is a desirable feature. Test work had been done with samples prior to the results discussed herein. This paper will discuss results of test work which were the basis for erection of a plant, and the subsequent plant operations. Laboratory Calcination and Air Tabling Tests Laboratory calcination tests showed that substantial upgrading of the borate could be accomplished by calcining followed by screening of the calcined material. Removal of the + 28 mesh calcine resulted in borate losses of less than 10% with a rejection of 40 weight % or more of the calcine. The minus 65 mesh calcine generally met the requirement of containing 48%, or more, B203. The minus 28 plus 65 mesh material contained an intermediate quantity of borate and would require additional treatment. Testing demonstrated that ore would not have to be reduced to a size finer than 19 mm (3/4 in.) prior to calcination. A temperature range from 400 to 455OC (750 to 850°F) was apparently satisfactory. Calcination at a temperature of 48Z°C (900°F), or higher, was unsatisfactory due to fusion. All laboratory calcination tests were static tests conducted by placing small covered charges in a laboratory furnace for 40 min. In all tests vapor issues from the furnace for 5 to 7 min. Following this period the ore could be heard "popping," due to decrepitation of the colemanite. The reaction generally continued for approximately one-half hour. Various size fractions of the calcination products from laboratory tests were subjected to laboratory air tabling tests, usually after removing the plus 28 mesh material. Laboratory air tabling tests were conducted employing a Whippet V-80 model air table manufactured by Sutton, Steele and Steele Co. now known as Tripple S Dynamics. Variables include both end and side-tilt, speed of vibration, and quantity of air rising through the deck. In addition to the variables in the machine itself, the feed rate is also a rather critical variable. Testing demonstrated that all - 28 mesh size fractions of the calcine could be successfully concentrated to 48% F2O3 or greater. For the finer material recoveries into the concentrate were between 85 and 90% of the borate. With the coarser material a substantial amount of middling was produced which required cleaner tabling. Laboratory calcination and air tabling tests indicated a process whereby the borate could be concentrated to about 50% B203 with borate recoveries approaching 90%. Moreover, the iron content of the concentrate was well below the required specification of 0.3% Fe2O3. Pilot Plant Calcination Following the laboratory test work described above, pilot plant testing was conducted to prove the process, provide data for engineering studies, and provide product for a prospective purchaser. The kiln used was 0.9 m-diam (3 ft) by 9.0 m (30 ft) long and had a belly section 1.2 m-diam (4 ft) by 2.74 m (9 ft) long near the discharge end. The kiln was operated at a speed of 0.7 rpm. Gas was fired into the kiln at an average rate of 27.1 m3/hr (958.4 cu ft per hr). The air to gas ratio used was 10:1. The ore was fed to the kiln countercurrent to the flame and discharged through a hopper into a screw conveyor which discharged to a 1.2 m (48 in.) Sweco separator. The separator had 28, 65, and 150 mesh screen cloths, with the plus 28 mesh fraction being discarded. The minus 28 mesh fractions were later subjected to air tabling. The exit gases, containing some calcine dust, were swept through two cyclones to recover the dust. The gases then were scrubbed in a Ducon scrubber; very little dust reported past the first cyclone. The dust from the first cyclone was also saved in drums. In addition to the gas rate, the flue gas velocity, after

Jan 1, 1981

By Kenneth E. Gray

The turbulent flow drag coefficients, or friction factors, have been experimentally determined for the cut-tings normally encountered in drilling operations. The gas law and average drag coefficients for characteristically flat particles (limestones and Shales) and for angular to sub-rounded particles (sandstones) were used to extend Newton's equation to a more useful form. The resulting equations are expressions for the slip velocity of a particle as a function of the particle size and shape, the bottom-hole injection Pressure, and the Bottom-hole temperature. The velocities necessary to lift actual rock cuttings as observed in the laboratory were compared with velocities predicted from the derived equations. Results indicate that there is no single correct circulating velocity and that the commonly used 3,000 ft/min linear velocity may he sufficient to lift only very mall particles. The advantage, in terms of horsepower requirements, of cycling high pressure air to obtain lower compression ratios was shown. INTRODUCTION A promising departure from standard rotary hydraulic drilling is air or gas drilling. In many instances the rate of penetration and bit life increases resulting from this method have been very substantial. Further the application of air drilling in areas of lost circulation or water sensitive pay zones has been highly successful. Purpose For the drilling contractor, the practical question remains, "How much air pressure and volume should I have?' lt is necessary to have a reasonable knowledge of the air velocities and pressures required in air drilling operations for the proper design and most advantageous use of surface equipment and for adequate removal of cuttings from the borehole. Eased on experience, most operators agree that at least 3,000 ft/min linear velocity is necessary for satisfactory hole cleaning.' It was the purpose of this investigation to evaluate the drag coefficients, or friction factors, for the cuttings normally encountered in drilling operations, i.e., sandstones, limestones and shales, and to utilize the values obtained to develop an expression for the minimum velocities and pressures (hence the volumes) necessary to lift these cuttings. THEORY Several investigators have done extensive work on the ability of drilling mud to lift bit cuttings. They conclude that the ability of a drilling mud to lift cuttings is dependent on the density difference between the rock being drilled and the drilling mud, the cutting size and shape, the mud flow constants, and the flow state (laminar, transitional or turbulent) of the mud. However, drilling muds are not true fluids but have a variation of viscosity with velocity in laminar flow and a constant apparent viscosity in turbulent flow. A survey of the literature revealed that no experimental work had been done to determine the velocities necessary to lift actual rock cuttings using a compressible fluid such as air or gas. In 1953, Nicolson6 stated that in air drilling fluid mechanics the particles could be assumed spherical in shape and that ; constant drag coefficient of 0.5 could be used due to the highly turbulent flow of the circulating air. By equating the turbulent resistance on the particle to the gravitational force on the particle, he obtained the expression, ^ = 2.67^y ,.......(1) where V, is slip velocity of spherical particle, ft/sec; D is particle diameter, in.; p, is particle density, 1b/ft8; and p, is fluid density, lb/ft3. However, as stated earlier, most operators seem to agree that adequate hole cleaning can be obtained by circulating a sufficient volume of air to give a linear velocity of 3,000 ft/rnin in the annular space. Basic Equation The velocity with which a solid particle freely falls through a fluid will increase until the accelerating force, gravity, is equal to the resisting forces, buoyancy due to the fluid displaced by the particle and friction due to the relative motion of the particle through the fluid. When the accelerating and resisting forces become equal,

By R. V. Higgins

This report presents the results of a study of water injection in the HX sub-pool fault block V-B, in the Upper Terminal zone, Wilmington field, California, and is based on data obtained to May 1957. An earlier report3/ describes the results of gas injection in part of the field. The principal purpose of the pilot flood was to determine whether a sand having a permeability of 1,200 millidarcys and containing oil having a viscosity of 17 centipoises could be waterflooded effectively. The results of a preliminary study by engineers of the Long Beach Harbor Department of oil recoveries from areas where encroaching edgewater was active and the results of an independent study by a consultant retained by the Long Beach Harbor Department were favorable to a trial waterflood. After these preliminary studies, a detailed petroleum-engineering exam-ination was made of the technical details. As a result, a peripheral flood was started in June 1953. The peripheral flood was deemed desirable because the producing wells are bottomed a short distance from the edgewater. The HX subpool was selected for the pilot test because it is an isolated section of the Upper Terminal zone and is separated from the adjoining pools by faults on the east and west and by edgewater on the north, south, and west. Several petroleum-engineering methods were applied to evaluate the performance of the waterflood, and all showed that injecting water stimulated oil production.

Jan 1, 1958

By M. A. Trevits

Mine seals are used in underground coal mines throughout the United States to isolate abandoned mining areas from the active mine workings. The objective of adding inert gas to a sealed mine area is to quickly reduce the oxygen concentration in the sealed area to a level that will not support combustion. In the US, this process is typically conducted from the ground surface and because surface-to-mine access is constrained by so many variables, it is very difficult to provide a source of inert gas to the place where it is needed underground. To alleviate this problem On Site Gas Systems, under a NIOSH contract, built a nitrogen (N2) generation plant that operates in the underground environment. The purpose of this work was to create a reliable in-mine mobile plant that would extract N2 gas from the mine atmosphere and use the gas to create and maintain a safe sealed mine area. The plant is based on a novel design using pressure swing adsorption technology and is sized to fit on a standard shield car or mine dolly for easy transport in and around an underground coal mine. The plant is capable of producing about 900 scf/min of N2 gas at ambient mine conditions. After an exhaustive series of laboratory tests, the plant was field tested at the NIOSH Safety Research Coal Mine (SRCM) where a portion of the mine was sealed and rendered inert. During the test, the movement of the injected gas was monitored at several underground points to observe the progress and extent of the inerting process. This paper describes the plant design and the results of the field test at the SRCM. Disclaimer: The findings and conclusions in this report are those of the authors and do not necessarily represent the views of NIOSH.

Jan 1, 2009