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By Frederic Benard

THE Ontario Refining Company, Limited, wholly-owned subsidiary of The International Nickel Company of Canada, refines the entire blister copper output of the latter's Copper Cliff smelter. Prior to 1936, the blister was cast in 460-lb. cakes at the smelter, loaded on open cars, and transported by rail to the refinery, a distance of approximately 1 1/4 miles. Here the cakes were transferred to the plant narrow-gauge system, charged to furnaces, and re-melted. This procedure necessitated the operation of holding furnaces and casting machines at the smelter -extensive re-handling-and melting down of a cold charge at the refinery. Transfer of the copper in molten form direct from converters to anode furnaces promised a considerable reduction in costs, and this possibility was investigated. As a result, a portable holding furnace, or hot-metal car, was designed and constructed. It was placed in operation in July, 1936, to be followed a year later by three additional units of slightly modified design. Figure 1 shows the original car, and Figure 2 the lacer type. With four cars in service, operations were placed on an all 'hot-metal' basis.

Jan 1, 1938

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION The two unit operations in cut-and-fill stoping that would benefit most from increased mechanization are drilling blastholes and moving broken ore to the ore¬passes. Drilling in current practice is done with hand¬held jackleg and stoper drills. Broken ore is commonly moved from the face to the orepass by a small slusher. This type of equipment can be easily moved through the raises. Hand-held drills have a slow drilling rate and require one operator for each machine. Drilling can be done much faster by small jumbo-mounted machines as two or more drills can be operated by one miner. Load-haul-dump (LHD) units move broken ore from the face to one or more orepasses at a much faster rate with less loss and dilution than slushers. These units are mobile and may be used at more than one place during a shift. The problem which must be overcome in applying this higher degree of mechanization is the restricted means of access inherent in current practices. Cut-and¬fill methods are usually applied to mining veins that are narrow, irregular in dip and strike, or where ground con¬ditions require the support of sand fill. The mechanized equipment which is needed to accomplish high produc¬tion cannot be raised or lowered through raises without being disassembled. A jumbo which will give the desired performance would be mounted on a rubber-tired carriage with a diesel engine for propulsion. Twin hydraulic booms run by a small air motor would carry pneumatic drills. Over¬all dimensions of the unit would be about 8 m (26 ft) in length, 1.6 m (5'/z ft) in width, and 1.6 m (5'/z ft) in height. An LHD of the desired capacity would be rubber¬tired and either diesel or electric-powered. The overall dimensions of a 2.3-m (3-cu yd) unit will be about 8 m (26 ft) in length, 1.8 m (6 ft) in width, and 1.5 m (5 ft) in height. It will weigh about 12 700 kg (28,000 1b). It can readily be seen that conventional raises can¬not handle this size equipment. The most desirable means of access enables equipment to be driven into the stoping areas. Ramps can enter the stope by being driven upward in the vein walls or by being developed in the vein on top of the fill as the stope advances. This ramp system not only permits the use of rubber-tired equipment in multiple working places but also enables its removal for repairs without component breakdown. In the proposed method, as in conventional cut-and¬fill stoping, sand fill will replace the mined-out ore. By constructing an inclined timber floor at one end of the stope, a ramp can be maintained on top of the fill and advanced as the stope is advanced cut by cut (Fig. 1). The ramp is, therefore, developed during the stoping process and is only driven as a tunnel when rising over and around the sill and crown pillars. A stoping section consists of three blocks of different geometric shapes. The first and last blocks to be mined will be triangular in the plane of the vein, and the blocks between will have the shape of parallelograms (Fig. 1). The first block will develop the initial ramp by breast stoping and filling to the ramp floor (Fig. 2). The ramp will need to be completed through to the next level be¬fore stoping of the subsequent block can begin. The subsequent blocks are mined by back stoping with the next ramp being developed at the other end of that block (Fig. 3). The length of each subsequent block can be varied without changing the slope of the ramp. As the new ramp is developed, the previous ramp is filled along with the stope cut. Access up and down is maintained through the stope. The final block is again breast stoped as no new ramp will be developed at the end line for jumbo egress. Boreholes provide ventilation between levels as well as utility access. They can also provide a secondary escapeway for personnel during periods of development. These boreholes would be located about 3 to 4.5 m (10 to 15 ft) in the footwall and spaced about 183 m (600 ft) apart along the vein. The orepasses are spaced about 61 m (200 ft) apart in the vein. This spacing allows not only a short tram for the LHD unit but also provides surge capacity for the ore between haulage trains. These passes are raised from the haulage levels. As development progresses, the raises are driven through the sill pillar, and then car¬ried in the fill as stoping progresses. Ventilation and utilities can be maintained through an orepass when it is not needed for ore. The end line orepasses are intended for this purpose as well as for personnel emergency egress. Overall dimensions and criteria assumed to illustrate the ramp method of access in a highly mechanized cut¬and-fill stope are: haulage level spacing of 61 m (200 ft); ramp gradient of 17%; average dip of approximately 1.57 rad (90°); initial stope length of 358 m (1177 ft); subsequent stope length of 179 m (588 ft); average stoping width of 3 m (10 ft) ; sill pillar thickness of 3.6 m (12 ft); crown pillar thickness of 1.8 m (6 ft); bored raises of 1.5-m (5-ft) diam; and cribbed orepasses of 1.2 x 1.2 m (4 x 4 ft) through the sill pillar and 1.2 m (4 ft) square inside the cribbed portion. DEVELOPMENT The ramp access system lends itself well to conduct¬ing the several phases of development concurrently. Haulage Drifts and Raises The first development step will be to drive the 3 x 3-m (10 x 10-ft) haulage drifts above and below

Jan 1, 1982

By B. W. Vallat

The Newport mine, located at Ironwood, gogebic county, Mich., on the Gogebic iron-range, is owned and operated by the Newport Mining Co., for the mining of iron-ore. I. GEOLOGY. The general geology of the gogebic range has been determined and recorded at different times by well-known geologists,' and their work will be referred to briefly in connection with the local conditions in the mine. The strike of the formation across this property is about 15" north of east. The general dip is 68" to the north. At the base of the formation on the aouth lies the granite. Looking north, at right angles to the strike of the formation, we have the foot-wall of quartz-slates and quartzite about 400 ft. wide; the iron-formation about 800 ft. wide, and the hanging-mall of black slates. The iron-formation is composed of banded jasper and quartzite together with iron oxide and concentrations of iron-ore. In character the ore iu a soft red hematite, with occasional masses of hard blue "steel ore." The formation is crossed in many places by diorite dikes, which cut through at various angles, the larger or main dikes dipping into the foot-wall in a SE. direction at angles varying from 15" to 30°, while numerous small dikes occur which strike through the formation in a vertical plane and at approximately right angles to the main dikes. There are various theories concerning these dikes, especially as to the time of their origin with relation to the ore-deposition. The commonly-accepted theory is that the dikes were there first, but there are some who contend that they were formed

Jan 1, 1912

By W. G. Scott

Since devising the incline top-slicing method in use at the Coronado mine,l I have had numerous inquiries as to how the same system could be adapted to larger orebodies. Based upon our experience here, I have outlined below some suggestions for its application to orebodies of larger size. Modifications and extensions, of course, would have to be made to take care of different dimensions and irregularities. In order that one may have a better understanding of the scheme, a brief description of the Coronado method will be given first. CORONADO METHOD The portion of the Coronado vein that is worked by the slicing system ranges from 20 to 40 ft. (6 to 12 m.) in width. Preparatory work in this orebody consists of driving sub-levels every 55 ft. (1 6.77 m.) vertically, or in multiples of 11 ft. (3.36 m.) (which is the height of a slice). These levels are usually driven along the hanging wall until the end of the orebody is reached. Crosscuts, at right angles to the strike, 50 ft. (15.24 m.) apart, are then run from wall to wall, or until the lean ore is reached. They are run in the middle of the block, which is 50 ft. in length and the width of the vein. The crosscut on the hanging-wall side is extended in the wall a short distance for a manway and timberway raise. This raise is carried up in the wall to the sub-level above, leaving a pillar about 5 ft. in thickness between it and the vein. Connections are made every 11

Jan 1, 1918

By C. Vasilescu

Phytotechnologies use plants to remediate various media impacted with different types of contaminants. These technologies can be implemented either in situ or ex situ. The affected media in which phytotechnologies can be applied include soil, sediments, groundwater and surface water. They are best applied at sites with shallow contamination of organic, nutrient, or metal pollutants that are amenable to one of five treatment options: phytoextraction, phytotrans-formation, rhizosphere, bioremediation, phyto-stabilization or rhizofiltration. Phytoremediation can be used along or, in some cases, in place of mechanical cleanup methods. Phytoremediation can be used to clean up metals, pesticides, solvents, explosives, crude oil, polycyclic aromatic hydrocarbons and landfill leachates.

Jan 1, 2006

By D. R. Nagaraj, Napoleon Tercero, Raymond Farinato

"Several studies conducted recently have shown that molybdenite flotation from Cu-Mo ores is adversely impacted in sea water at pH >~ 10.0 irrespective of the pH modifier; under the same conditions Cu sulfide flotation is relatively unaffected. This has been attributed to the presence of Mg2+ ions and their hydrolysis products in the sea water which adsorb and precipitate on molybdenite surfaces as Mg(OH)2. However, why and how Mg(OH)2 impacts only molybdenite flotation in the complex flotation pulps containing a multitude of minerals is not understood. The work presented in this paper probes the interface and interphase regions near the edge and basal planes of natural molybdenite (MoS2) in aqueous solutions and flotation pulps, particularly in the presence of copper ore pulp and added Mg2+. Differences in interaction of metal ions and their hydrolysis products (Mg2+, Mg(OH)+) were measured on the two faces of this anisotropic mineral. How those differences relate to molybdenite floatability and the impact of aqueous solution chemistry are discussed.INTRODUCTIONWater resource management is an important sustainability challenge in mineral processing. In view of the shortage of fresh water in many parts of the world, mineral processing plants are evaluating alternative sources. Seawater flotation has received much attention recently and there have been several important studies reported in the literature, particularly for the processing of Cu-Mo ores (S. Castro, 2012; S. Castro, Rioseco & Laskowski, 2012; S. Castro et al., 2012; Laskowski & Castro, 2012; Ramos, Castro & Laskowski, 2013; S. Castro & Laskowski, 2013; Laskowski, Castro & Ramos, 2013; Nagaraj & Farinato, 2014; S. Castro, Uribe, & Laskowski, 2014; Veki, 2013). The main conclusions from these studies on Cu-Mo ores in seawater at elevated pH were: a) flotation recovery of Cu sulfides is either unaffected or only marginally reduced, b) molybdenite recovery is adversely affected in seawater at pH values necessary for rejecting pyrite (i.e. pH > 10 to 10.8 depending upon particular system), and c) the precipitation of Mg2+ in seawater as Mg(OH)2 is responsible for the adverse effect on Mo flotation. In order to further investigate the individual effects of Mg2+ and Ca2+ on molybdenite depression in Cu-Mo ore flotation, Nagaraj & Farinato (2014) conducted flotation tests wherein individual metal ions were spiked into tap water during grinding of Cu-Mo ores. They showed the strong adverse effect of Mg2+ addition on Mo recovery and, to a lesser extent, on Cu recovery when pH > 9.5. Under similar conditions, Ca2+ had virtually no effect on either Cu or Mo sulfide flotation. Figure 1 below shows the previous results for Cu- Mo ores in tap water and the effect of addition of Ca2+ and Mg2+ ions. Note that in seawater the adverse effect of Mg2+ on Cu recovery was much smaller than that in tap water (Nagaraj & Farinato, 2014; Figure 4a and 4b and Table 4 therein)."

Jan 1, 2016

OFFICERS OF THE WOMAN'S AUXILIARY AIME President-Mrs Robert D Coombs, 375 Paramus Road, Paramus, N J First Vice-President-Mrs W McG Peirce, Troy Court 73 F, New England Avenue, Sum¬mit, N J Second Vice-President-Mrs Andrew E Beer, Richardson Avenue, Sea Cliff, Long Island, N Y Third Vice-President-Mrs Ira T Hook, 494 Norton Parkway, New Haven, Conn Fourth Vice-President-Mrs John R Kenney, 427 W Wrightwood Avenue, Chicago 14, Ill Fifth Vice-President-Mrs Alfonso Ballon, Montero Rosas 115, Barranco, Lima, Peru Recording Secretary-Mrs Arnold H Miller, 601 West 115th St, New York 25, N Y Assistant Recording Secretary-Mrs R J Mechin, 66 East 79th Street, New York 21, N Y Corresponding Secretary-Mrs Theodore Nagel, 76 Remsen St, Brooklyn 1, N Y Treasurer-Mrs Robert W Rowen, 116 East 68th St, New York 21, N Y Assistant Treasurer-Mrs Samuel H Dolbear, 11 Broadway, New York 4, N Y Assistant Treasurer-Mrs H A Kursell, 12 Murray Hill Road, Scarsdale, N Y

Jan 1, 1956

By Richard N. Appling

This report presents information obtained by examination and investigation of manganese deposits in northeastern Oregon during May and June 1954 and April 1957. Maps, assays, and other examination data are presented on 10 manganese deposits -9 in Baker County and 1 in Grant County. Included among these deposits are all those that have undergone substantial exploration or have a record of production. Seven of the Baker County deposits occur as characteristically small, irregular pods and lenses in argillite. The remaining Baker County prospects are [marily] narrow, vein-type occurrences in schist. The Grant County occurrence in greenstone and serpentine. All the deposits are composed of various manganese oxides, intermixed with abundant quartz and chert.

Jan 1, 1959

By Gregg Wilkerson

The Blind Springs Hill pumice beds are located 5 miles southwest of Benton, California in the foothills between the Benton Mountain Range and Blind Springs Valley. The best exposures of the pumice are in an area 0.75 miles west of Blind Springs Hill. The pumice is Quaternary ejecta material which came from the rhyolitic volcanoes in the Glass Mountain-Volcanic Tablelands area of Mono County. The pumice beds occur in two distinct layers separated by an unconformity. The upper layer is 0 to 11 feet thick and consists of coarse subangular pumice fragments, some of which are over 1.75 inches in diameter. The lower bed is medium to coarse grained with subangular fragments up to 1.0 inch. The lower bed is 0 to 9 feet thick. Overburden at the Blind Springs Hill deposit varies from 0 to 15 feet. There are approximately 1 million cubic yards of material in the deposit that can be mined. The size distribution in the deposit is mostly fine with only 10.6% of the pumice fragments being over 0.5 inch in diameter, the minimum size needed for stone-wash or acid-wash applications. Suitability of pumice for the textile industry is quantified by standard 15-minute abrasion and absorption tests. In the Blind Springs Hill deposit, the plus 0.5 inch fraction has an abrasion loss of 33% which is too high for stone-wash application but it has an absorption capacity of 36% which makes it suitable for acid-wash applications in the textile industry. The abrasion and absorption properties of this deposit are within the range of those which have been measured for other pumice deposits within the United States.

Jan 1, 1995

By Griffith L, Lama RD

West Cliff Colliery mines Bulli Seam occurring at a depth of approximately 480 m. The Seam is intersected by a number of small shear zones and faults. While the faults have caused displacement of the seam, the shear zones are strike-slip structures with small apparent displacement of seam. Outbursts at West Cliff Colliery occur while driving through these shear zones. Few outburst wave occurred where faults have been intersected. In the last 3 years, over one hundred outbursts have occurred at West Cliff Colliery. The size of outbursts varies from a small bump to exceeding 140 tonnes of coal. A number of studies have been undertaken at West Cliff Colliery to control outbursts and predict Outbursts in advance of workings. These include: geological investigations, microseismic techniques, gas adsorption studies and gas drainage. The investigations to date suggest that outbursts of gas at West Cliff Colliery are more a gas phenomenon similar to what occurs in overseas mines. The results of various investigations are reported. West Cliff Colliery is located in the mid- west of the southern coal fields of N.S.W.

Jan 1, 1980

By Earl Oliver

Page A.I.M.E. Stabilization Committee Activity................. 133 Definition of Capture Rule: Robert E. Hardwicke.......................... 134 Westmoreland Natural Gas Co. vs. DeWitt................ 135 Kelly vs. Ohio Oil Co......................... 135 Bernard vs. Monongahela Gas Co.................... 136 Definition of Ownership in Place..................... 137 Prospective Stabilization Remedies.................... 137 Dean Roscoe Pound's Four Teats of Practicability.............. 137 I. Mechanically Feasible....................... 138 Capture Rule.......................... 138 Ownership in Place....................... 140 Mechanical Separatio.7 Method................. 140 Engineering Appraisal Method................. 141 II. Economically Advisable...................... 143 Capture Rule.......................... 144 Ownership in Place....................... 144 III. Politically Wise.......................... 144 Capture Rule..........................144 Ownership in Place.......................147 IV. Legally Sound..........................148 Capture Rule..........................148 Ownership in Place.......................149 Dean Henry M. Bates' Analysis......................149 Raumé................................152 The A.I.M.E. Stabilization Committee was named in 1932 to ascertain whether some practical substitute could be devised for the Rule of Capture, which was generally credited with being the major cause of instability in the Petroleum Industry. After considerable intercommittee correspondence a letter was received from E. DeGolyer under date of Jan. 25. 1935. in which he discussed the possibility of modifying rather than discarding the Rule of Capture. He summarized his proposal as follows:

Jan 1, 1937

By P. F. Ziemkiewicz, A. S. Knox

Large-scale use of CCBs in mine environments involves filling surface and underground coal mines with ash to address hydraulic problems, AMD, backfilling and subsidence. The environmental risk of land filling CCBs is generally evaluated by the Toxic Characteristics Leaching Procedure (TCLP) or the Synthetic Precipitation Leaching Procedure (SPLP). However, there is doubt regarding the applicability of these tests to long-term CCBs leaching behavior in groundwater associated with coal mines. The Mine Water Leaching Procedure (MWLP) was developed to provide a site specific risk assessment tool. The MWL Pprocedure is presented in this paper and compared with TCLP as a study.

Jan 1, 2003

By Donald C. Craighead

"Overland conveyor systems to transport bulk materials from mines to processing plants or load-out stations have to be seriously considered as the viable alternative to truck and train haulage systems. With the ever-increasing price of diesel fuel, conveyors with their electrical drives generally prove to be the least costly when operating and maintenance expenses are analyzed for the life of the installation.IntroductionOverland conveyor systems are now considered as one of the viable alternatives which should be closely investigated when a mining company and its engineers initiate studies on methods of transporting raw materials from a mine to a processing plant or rail load-out station.This paper will generalize the overland conveyor and compare it with alternative handling systems. Each installation must be carefully studied to determine the best route, type of foundations river crossings; amount of excavation work for cut and fill, maintenance roads, etc. No two locations are the same nor are the principles of operation of an overland conveyor the same as those of a normal plant or mine belt conveyor. The proper time and effort must be put into the study of the system to evaluate the costs for comparison with other systems.The length of the overland conveyor depends on the layout of the mine and plant and the topography of the terrain between the terminals. Individual lengths may be up to 20 km, and a series of conveyors may extend 100 km or more.Advantages of Overland ConveyorsThe main advantage of overland conveyors is the low operating and maintenance costs which will be realized over its years of operation. Although truck haulage probably will permit lower capital investment during the initial phases of the mine development when the annual capacities have not developed to the design level, their operating and maintenance costs will be considerably higher. When evaluating the total costs per ton of material to be handled during the expected life of the operation and the projected increases for diesel fuel versus electricity, overland conveyor systems will generally prove to be economically more viable."

Jan 1, 1982

Effective strata control, utilising fully encapsulated roof bolts, is dependent on the installed quality of the reinforcement elements. One mechanism by which fully encapsulated roof bolts may become less than fully effective is by gloving (glove fingering) and un-mixing of the resin. Following some small roof failures containing gloved bolts, investigations have been undertaken in some New Zealand coal mines to determine the extent of, and mechanisms involved in, the gloving of fully encapsulated roof bolts for a range of roof types and installation methods. Gloving, in this context, refers to the resin cartridge partially or completely enveloping a length of the bolt, typically with unmixed resin filler and catalyst/hardener remaining within the cartridge. Results have shown that gloving and un-mixing is a systematic and widespread phenomena, occurring across a range of resin and/or bolt manufacturers, and across a variety of roof types. Gloving was found in bolts installed using: hand held pneumatic and continuous miner mounted hydraulic methods, under run of mine (ROM) conditions by mine workers, and under controlled "best practice" conditions. The roofs bolts investigated are all 1800mm (6ft) long, and have shown that, on average, 485mm (-19") of the bolt length is gloved and/or unmixed. This equates to a 27 percent reduction in effective bolt length. The effected length of bolt is sufficient to impact on the load transfer/reinforcement capabilities of the bolt system, and hence impacts on the design assumptions and subsequent stability assessments. Furthermore, there is a direct relationship between the mechanism for the gloving of bolts and an increased occurrence of resin loss and reduction in encapsulations length in soft (less than 10 MPa) and/or highly fractured rock. Investigations have concentrated on developing an understanding of the mechanisms involved. This has been undertaken using both field installations and recovered bolts, as well as test bench trials. The mechanisms involved have been confirmed as being: the development of a pressure front as the bolt encounters the resin and is spun up the hole, which, in turn, leads to over-pressurisation and radial expansion of the resin cartridge, resulting in an increase in diameter of the plastic resin cartridge, which allows the bolt to be spun up into the cartridge without making sufficient contact to shred the cartridge or the hardener tube, resulting in an unmixed portion of resin.

Jan 1, 2003

By V. Schenk

Dust exposure resulting from the exploitation and winning of natural resources below ground has presented a serious health problem since the beginning of mining. The boom phase of the German hard coal mining industry after World War II with at times up to 500,000 employees saw an increase of dust exposure of miners below ground as the mechanization of road heading and mining increased significantly. More and more miners were diagnosed with coal-workers-pneumoconiosis (CWP), a chronic and irreversible lung disease resulting in a significant impairment of pulmonary function. This realization led to a reduction of CWP in the German hard coal mining industry in the 1960s. The successful fight against CWP began with the introduction of technical dust prevention measures. By infusing the coal seam with water before mining, sprinkling and spraying of heading and mining machines as well as canopies with water, aspiration in places with high amounts of dust and the suppression of dust with chemical substances it is possible to significantly reduce the miners? dust exposure. All operating points below ground are checked and measured regularly with regard to dust exposure. Through the combination of these values together with the individual shift data a relatively precise profile of individual dust exposure can be obtained. This profile allows the targeted work assignment when certain predetermined limits have been reached, moving the miner to a work place with less dust exposure and thus reducing the chronic dust exposure. The state of health of the individual is known because since about 60 years ago extensive occupational health checks have been carried out regularly at least every two years which means that work restrictions can be imposed depending on the findings. Since the beginning of the 1980s more value has been attached to not only situational but also behavioral prevention. Given the comprehensive provision of protective masks, having a protective effect as well as the required comfort when being worn, each miner has the opportunity to protect his respiratory organs depending on the specific dust exposure. Through a combination of all these measures the German hard coal mining industry has succeeded in beating the classic coal miner?s disease CWP. For mines in the Ruhr valley this means that a miner who started work after 1978 will no longer develop CWP.

Aug 1, 2013

By S. B. Breyner, A. Ressler, D. Gregurek, A. Spanring

"The changes in process conditions that frequently decrease the service life of refractory products provided incentive for RHI AG to develop high performance bricks, especially for non-ferrous metal applications. Innovative product development was required to determine suitable refractory products for improving lining life. At the Technology Center in Leoben, Austria, practical corrosion testing methods and facilities, including an induction furnace and rotary kiln, enable detailed comparative analysis of brick wear on a pilot scale. In tests that were carried out using a fayalite type slag, the newly developed RADEX VFG magnesia-chromite grade showed very promising results compared to the standard grades. Based on the test results, optimal refractory solutions can be recommended for this challenging industrial application and field trials can be performed directly at the customer's site.IntroductionThe complexity of metallurgical processes in the non-ferrous metal industry, and especially changes to the process conditions in combination with the diversity of metal processing furnaces ''', requires a precise understanding of each individual system for appropriate refractory product selection l2'. In addition, particular importance is focused on the specific refractory consumption being as low as possible.Another important aspect results from the global shortage of resources, which forces companies to invest in recycling processes. In the base metal industry, recycling operations have resulted in significant changes; for example, operating with two slag chemistries and process temperatures will influence the performance of previously suitable linings. The raw material mix for primary smelters has also changed, including higher levels of secondary raw materials in the smelter feed."

Jan 1, 2012

In the course of an investigation of the influence of bismuth on copper carried out some years ago, (1) the author found that phosphorus exerted a beneficial effect on bismuth-copper alloys of low bismuth and oxygen content by neutralizing the embrittling effect of the bismuth. In the intervening years, a further examination of bismuth-phosphorus-copper alloys has been canied out and the influence of other elements and combinations of elements on copper containing bismuth has been studied.The results recorded in the present paper are part of this investigation.MATERIALSAll the alloys were made from high-grade copper (KS.A. wire-bar) of the following composition: copper 99·95%, oxygen 0,040'%, iron 0·0015%, sulphur less than 0-001%, bismuth less than (j>'OOl%, other elements each less than O· 0010/<-. The bar was hot rolled and cold drawn to strip which was cut into short lengths and cleaned in nitric acid ready for melting.A piece of this 'strip was drawn to wire 0·08 in. diam., annealed at 5000 C, and measured for electrical conductivity as a check on purity. The result was 101·5% of the I.E'.C. standard.In making the alloys, phosphorus was added as 14% phosphor-copper, bismuth as 17% bismuth-copper, and the other constituents either as copper alloys or as elements of a high degree of purity.PREPARATIQN OF STRIPSThe charges were melted in plumbago crucibles (Morgan salamander brand) in an oil-fired furnace, Le;, in the manner commonly adopted in industrial practice. The copper was first melted under charcoal, well stirred with a chaned stick, and the necessary alloying elements added and stirred in with the stick. The weight of a charge varied according to requirements but generally was about 4 lb. The charges were poured into slightly tapered brass moulds heated to about 1000 C and dressed with a mixture of mineral oil and tallow. They were not only poured under charcoal through the burning mould dressing but also- through a ring of flame produced by burning kerosene on asbestos cord wrapped round the top of the mould. Experiments carried out with plain copper proved that by this means oxygen could be excluded from the stream of metal. The billets were 7·5 in. long x O·83 in. diam. x 1·3 lb. weight.The billets were cropped about 1 in. from the top, cleaned in nitric acid, and measured for density by the usual method of weighing in air and -in water. As these measurements were required only as a check on casting, technique a high degree of accuracy was not essential (accuracy = -+ 0·2%).

Jan 1, 1942

By Bertha Farfán del Carpio Vera

En este resumen se presenta los estudios iniciales que nos permitirán descubrir nuevas propiedades de estos 2 elementos, al ser transformados en áridos livianos, cuando luego de un proceso de precipitación química y a través de la catálisis heterogénea, son depositados en una sustancia de soporte conocida como zeolita clinoptilolita. Ambos metales Pb y Zn, adhieren totalmente a la zeolita, la cual se comporta como un catalizador metálico dispersado, siendo los metales Pb y Zn, los que se depositan sobre la anterior pasta para formar los pellets.Luego de mezclarse completamente, el soporte y el material catalítico se compactan, se someten a un proceso de extrusión.

Sep 1, 2017

By L. E. Mannion

Sodium carbonate (soda ash) is one of two principal commercial alkalis. Its principal competitor is sodium hydroxide. The use of sodium carbonate is recorded in ancient Egypt, where naturally occurring brines and solid salts provided impure soda for early glassmaking. Until the 19th century, soda ash (as well as potash) was recovered mainly by leaching wood or other plant ashes. In 18th-century Europe, for example, leaching seaweed ash with hot water produced a brown laundry lye. One of the better grades of this material was a Spanish product, barilla, which analyzed 24 to 30% Na2CO3 (Hou, 1942). The modern chemical industry was born in 1691 with the LeBlanc process for making sodium carbonate from salt, sulfuric acid, and lime. English industry used this process extensively in the 17th and 18th centuries. The next major advance was in 1863, when Ernst Solvay devised the ammonia-soda, or Solvay process. The first of many Solvay plants was built in 1874. These supplanted the LeBlanc method and dominated worldwide soda ash production until 1974, when mined sodium carbonate became predominant. Although considerably less important until recently, natural sodium carbonate deposits historically have supported small soda ash operations throughout the world. Shortages during World War II stimulated production of "natural" soda ash from such deposits, particularly at Searles Lake, CA and Lake Magadi, Kenya. More spectacularly, increased production from the vast trona deposits in Wyoming has virtually supplanted the Solvay process in the US. Principal producers of natural soda ash as well as the character of the source deposits are listed in Table 1. [ ]

Jan 1, 1983

By Gordon A. Gross

Banded siliceous manganese-iron bearing sediments recognized as distinctive facies of Algoma-type iron formation constitute low-grade resources of manganese in many countries. Manganiferous facies associated with oxide and carbonate facies of iron formation have iron to manganese ratios ranging from 0.2 to 2 and commonly about one. Manganiferous facies up to 30 m thick occurring separately or with other iron formation facies are associated with cherts, carbonates, shales, turbidites, tuffs, and lavas. Primary cherty manganiferous facies and their gondite metamorphic equivalents represent significant manganese resources and are protores for higher-grade deposits derived by secondary enrichment processes.

Jan 1, 1983