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By Victor F. Hollister

Introduction Trends of successful discovery technology and psychology emerge from a thoughtful study of various companies' exploration programs. These subjective patterns can be generalized into groups of factors for discussion. Here, for convenience, factors that influence exploration success are separated into management and geologic concerns. Most successful discoveries have a demonstrably favorable management combined with a technically competent geologic input. Both are required. But the management input is decisive. Management factors include the correct definition of management's role; cost of exploration; time needed for exploration; commodity selection; and the fostering of autonomy, creativity, and confidence in the geology department. Geologic factors include imaginative technical competence (field as well as academic), correct development and use of empirical exploration or prospecting models, and dedication. Success comes to those entities that are properly managed and have technological competence. A management team that directs a company's business has been compared to a dynasty. The dynasty lays out policy and follows up with appropriate action. In that management's power is pervasive, the term is appropriate. Management's role is, however, more important. Many examples can be cited where the geologic input was impeccable. But management somehow snatched defeat from the jaws of victory, and the exploration success was turned to a third party's profit. Management Factors Management's role can be broadly generalized to four parts. These are to identify the commodities it seeks, define its part in the exploration process, develop a functioning exploration department, and have the desire to succeed. It must have confidence in its geologists and delegate to them the authority and responsibility they need to operate successfully. Head office management cannot do the thinking for a field staff thousands of kilometers away. There is no substitute for confidence and delegation of authority. Successes came to those who had the freedom to act. Motivation is a management requisite. Management's role is to prevent destructive, internecine political battles within the corporate structure. Most large US mining corporations have well known histories of political strife that have damaged their exploration efforts. Management's role must, above all, include the desire to succeed. There are cases where management has been confronted with a commercial exploration success, but their reaction is to request additional study or exploration. Perhaps the problem of bringing a commercial find into production is beyond their capabilities. But a bona fide discovery sent back for additional work commonly is lost due to that management. The cost of exploration is a factor that management must accept. A variety of studies made during the past 40 years shows the cost per discovery to be $25 million, $38 million, $200 million, or even $290 million. These figures are derived by dividing the total exploration expenditure by the number of discoveries. However, successful management dynasties have a much lower cost per discovery during their management regime. Asarco, from 1948 through 1963, had one of the lowest cost discovery rates in the world. Duval Corp., from 1948 through 1968, had comparably low discovery costs. Placer Development, from 1955, was equally successful. Western Mining in Australia has had one of the longest lived dynasties associated with exploration success. Time is vital in exploration success. Most successful exploration case histories stress sufficient time in project exploration so that the target is well tested. A large target cannot always be adequately evaluated by a few holes (strat or concept testing). Serendipity helps in exploring large targets, but it best operates in a lengthy program. Numerous case histories include pleasant surprises. But these came mostly after the 10th hole. Commodity markets are important factors governing exploration success. Management's decision to explore for a specific commodity can doom its effort if that product suffers from severe oversupply. Unquestionably, progressive copper price increases from 1948 through 1968 helped convert marginal porphyry copper discoveries

Jan 8, 1985

By M R. Stevens

The Kauri Block (Hauraki Hill) andMatawai Valley epithermal vein systems, near Coromandel township, are hosted by hydrothermally altered Miocene andesites and Mesozoic greywackes of the Coromandel VolcanicZone. The hydrological setting of both geothermal systems was probably that of an andesitic volcanic arc. Two types of veins are recognised: 1. Simple veins composed of fine comb quartz, and 2. Complex (breccia) veins composed of vein fragments of sulphide-bearing, chalcedonic silica cemented by saccharoidal and fine comb quartz.   Pyrite, arsenopyrite, marcasite, and gold form the ore minerals. Both deposits historically exploited hypogene and supergene bonanza gold.   At Hauraki Hill, breccia veins up to 7m wide contain abundant fragments of recrystallised colloform banded-crustiform textured quartz veins ('Golden Cross' style) and were probably major channels for ascending deep mineralising fluids. Alteration assemblages are zoned with the higher elevations of Hauraki Hill-Bunkers Hill characterised by strong kaolinite alteration whereas deeper levels are dominated by illitic clays.   At Matawai, complex veins and hydrothermal breccias pinch and swell, and inflect, within a total width of veining and alteration that exceeds 50 m. The greywacke host rocks are altered to an assemblage of illitesmectite, quartz, chlorite and pyrite although there are pockets of kaolinite.   The proposed genesis of veins and ores in both deposits invokes episodic rapid mixing (quenching) of hot mineralising reservoir fluids with cool acid condensate waters, whilst acidification and increased fO2 triggered gold deposition. These hydrothermal systems heated up during their later stages and at Matawai visible gold was deposited from a late mesothermal stage. The main conclusion is that the northwest CVZ has excellent potential for large gold deposits (= 30 tonnes contained gold).

Jan 1, 1993

By J. W. Miller

This report deals with upgrading of the coal reserves of Raleigh County, VI. Va., to metallurgical standards. The data on samples herein correctly show the characteristics of the bed at the point where sampled and are used to indicate the general quality of the bed in the county. The principal results and conclusions are: 1. Raleigh County contains large reserves of high-quality, high-, medium-, and low-volatile coals. 2. The principal beds from which an increased amount of coal is expected to be mined in the future are: The Eagle, Powellton, No. 5 Block, Hernshaw, Fire Creek, and Pocahontas No.6.

Jan 1, 1954

By J. Bouchard, M. Levesque, D. Georges-Filteau, P. Radziszewski, G. LeBlanc

"Grinding mills are infamous for their extremely low energy efficiency. This paper reports the results of an investigation that characterized energy outputs in industrial grinding mill circuits. It identifies and discusses the potential for recovering energy from grinding circuits. Agnico-Eagle Goldex Division, Mine Canadian Malartic, and New Gold New Afton Mine participated in the study by providing operating data for three semi-autogenous grinding mills and four ball mills. Results showed that on average, 79% of the supplied electrical energy was converted to heat absorbed by the slurry, 8% was lost through the drive system, and approximately 2% was transmitted to the ambient air. Only 9% of the input energy was actually used for grinding. The remaining 91%—wasted as heat—could potentially be recovered using suitable technologies or integrated energy management systems. This topic remains to be further investigated. ¦ KEYWORDS Comminution processes, Energy recovery potential, Grinding mills RÉSUMÉ Les broyeurs sont tristement célèbres pour leur rendement énergétique médiocre. Cet article présente les résultats d’une étude qui caractérise les rendements énergétiques des circuits industriels de broyage et examine les possibilités de récupération d’énergie dans les circuits de broyage. La mine Goldex d’Agnico Eagle, Mine Canadian Malartic et la mine New Afton de New Gold ont participé à l’étude en fournissant des données de fonctionnement pour trois broyeurs semi-autogènes et quatre broyeurs à boulets. Les résultats montrent qu’en moyenne 79 % de l’énergie électrique fournie sont convertis en chaleur absorbée par la pulpe, 8 % sont perdus dans le système d’entraînement et environ 2 % sont dispersés dans l’air ambiant. Seuls 9 % de l’énergie primaire sont, de fait, utilisés pour le broyage ; les 91 % restants (perdus sous forme de chaleur) peuvent potentiellement être récupérés à l’aide de technologies adaptées ou intégrées dans des systèmes de gestion de l’énergie. Cette question demeure toutefois ouverte. ¦ MOTS CLÉS broyeurs, potentiel de récupération de l’énergie, procédés de comminution INTRODUCTION Grinding is largely recognized as a very inefficient process; energy efficiency estimates range from < 1 to 2% (Fuerstenau & Abouzeid, 2002; Tromans & Meech, 2002, 2004) when comparing the input energy to that required for generating new mineral surfaces. Criticizing an “ill definition of the reference for the output energy,” Fuerstenau and Abouzeid (2002) proposed a “more meaningful baseline” would be “the energy for producing new surface area by the ¬mens.” Based on this reference, they concluded that “the ball mill is reasonably efficient energetically,” exhibiting an efficiency of approximately 15% for quartz."

Jan 1, 2019

By R. A. Pallanch

This paper, which describes the milling practice at the Midvale concentrator, is one of a series being prepared by the Bureau of Mines. The Midvale concentrator, a custom lead-zinc selective flotation plant, is in Salt Lake County, Utah, on a flat site several hundred yards east of the Jordan River and about 200 yards south of the Denver & Rio Grande Western Railroad tracks leading to Bingham Canyon. The site is at an elevation of about 50 feet above the river, the highest available location, thus allowing maximum storage of tailings.

Jan 1, 1931

By S H. Begg

"*This is an abstract only*Most technical and economic analyses are performed to reduce uncertainty to help make decisions. Decision-making is about ranking alternatives - rarely are precise predictions required, if even possible. This simple observation has profound implications for modelling approaches, how they deal with uncertainty and the reliability of information provided by subject matter experts.CITATION:Begg, S H, 2013. Some reflections on uncertainty, decisions, models and people, in Proceedings The Second AusIMM International Geometallurgy Conference (GeoMet) 2013, pp 3-6 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Sep 29, 2013

By A Marton

Concerns with crest losses have seen changes to a number of design and operating parameters to retain the integrity of the wall and the berms of the Kidston Gold MinesÆ Eldridge Pit. Until trials began in 1999 most final wall designs utilised large diameter blastholes up to the presplit row to achieve the wall angle. The berm crests were being lost, wall support requirements were increasing and affecting mining schedules and several fault areas were putting pit access at risk. Standard designs had been determined during the early mining years and while periodic modifications had been made there was a need for major blast pattern revision. This paper traces the development of wall control blasting programs over the years, relating designs to the geology and the evidence in the walls with the theories of the day and includes the recent trials that were conducted in what is a most unforgiving geology. Trials and improvements to ensure the walls were safe were conducted even as the mine neared its economic life. Instrumented monitoring of trials blasts has greatly increased the understanding of the damage criteria and enabled rapid revision of designs. Recent changes included reduction in presplit energy, charge redistribution and use of smaller blast holes for the trim patterns. Determination of a minimum proximity-to-final wall limit for production blastholes, establishment of the correct standoff to berm crests and walls and selective use of presplitting techniques for protecting crests have resulted in cost effectiveness, improved bench crest retention and an overall reduction in visible wall damage.

Jan 1, 2001

By E. C. I. Hammerbeck

"Geological controls and constraints in space and time define the milieu of the nickel deposits of Southern Africa. Primary, magmatic deposits are of two types, e.g. synvolcanic or syntectonic deposits confined to certain orogenic domains, and those related to intracratonic, post-orogenic magmatism, e.g. the Great Dyke, the Bushveld Complex and the Insizwa intrusion. Epigenetic and supergene deposits form a separate, unrelated group and are of minor economic importance.The metallogenic evolution of syntectonic nickel deposits is particularly well illustrated in the Zimbabwean Province, where the relationships between Archaean greenstone development and nickel mineralization are best illustrated. Volumetrically subordinate, nickel-poor greenstones in the 3500 m.y. old Sebakwian Group are followed by more common komatiite development in the Lower Greenstones (3000 m.y.), and, finally, by very significant komatiites with important nickel mineralization in the 2700 m.y. old Upper Greenstones of the Bulawayan Group. Comparable nickel mineralization is not known in the Raapvaal Province, in spite of ubiquitous komatiitic magmatism of the Onverwacht Group (3500 m.y.) and other greenstone sequences. No komatiitic rocks younger than about 3200 m.y. are known in this region, orogenic activity in the Kaapvaal tectonic province having ceased at about 3100 m.y., whereas the Zimbabwe Province remained tectonically active until at least 2600 m.y.The tectonic environment, host-rock petrology and mode of emplacement, as well as some details of the nickel mineralization, are the parameters employed to classify the Southern African magmatic nickel deposits. A rating of the deposits is introduced that reflects the original resources of the deposit, its average nickel content, its mining history and potential to facilitate ready comparison with and assessment of the nickel deposits listed in the classification, as well as some important nickel deposits worldwide."

Jan 1, 1984

By C. L. Schneider, C. O. Gomes, R. P. King, J. A. Finch, D. Lin

Discussion is presented by Prof. King and Dr. Schneider of the paper by Lin, Gomez and Finch published in Trans.IMM C, vol.104, 1995, p.C155-C161. They conclude that the correction procedures used for simulated observations of simulated capped spheres are not reliable when applied to mineralogical data and that their relative ranking should not be extrapolated to other situations. Properly regularised solutions obtained with the use of appropriate kernels can be safely used for routine mineral liberation work and can be calibrated accurately against real mineralogical samples in the laboratory

Jun 19, 1905

By A. O. Ipsen

As part of the Bureau&apos;s intensive investigation of domestic manganese deposits, lots of ore for testing were collected by Bureau engineers during examinations of several properties in Juab County, Utah. One low-grade and two intermediate-grade samples (13.5, 25.3 and 27.5 percent manganese) were procured by engineers of the Bureau of Mines from the 900--and 1,100-foot levels of&apos; the Tintic Standard mine, situated 2 miles southeast of Eureka, Utah. The ore bodies were described as blocks of ore approximately 60 feet long, 40 feet wide, and 20 to 100 feet deep. The manganese-iron association in the Tintic Standard ores was intimate, and no significant concentration of the low-grade sample was made by ore dressing, However, both of&apos; the higher-grade samples were readily beneficiated to marketable manganese products. Ore from the Oxen No. 5974 claim, 2 miles northeast of Eureka, Utah was low in grade (19.4 percent manganese) and was not amenable to the production of high-grade concentrates owing to the intimacy of association of the manganese minerals with silica. However, a fair manganese recovery was obtained in a product that, when sintered, assayed plus 40 percent manganese. The ore body is reported to be a replacement in limestone and is composed of manganese oxides intermixed with silica, decomposed rhyolite, and limestone. The Fields property consists of five mineral claims situated 3 miles north of Eureka, Utah. The deposit is a replacement in a contact between limestone and a quartz dike. Tests were made on representative samples of a 50-ton lot stored on the property. Although the mineral association in the ore (22.2 percent manganese) is not particularly complex, the physical properties of the manganese minerals render the ore refractory to ore-dressing beneficiation. As the ores varied widely in physical and chemical characteristics, samples from each property were considered as separate ore-dressing problems.

Jan 1, 1949

By Brij M. Moudgil

Florida produced more than eighty percent of domestic phosphate during 1974. This amounted to more than thirty-five million short tons. Ninety percent of this production was used in the production of fertilizers. Other uses of phosphate include food preservatives, fabric dyes, optical glass, photographic films, soaps, plastics, pharmaceuticals and animal feed ingredients. Marketable phosphate rock projections for the nest decade forecast a very healthy growth for the Florida phosphate industry. Associated with the expansion, however, are the corresponding problems of water conservation, air and water pollution and land reclamation. Recognizing mining as an interim land use, the industry is presently engaged in a very active reclamation program. The proximity of the central Florida deposits to rapidly developing urban areas makes it important to undertake successful and well planned reclamation programs. Water use and conservation is of vitil importance to the industry. Water withdrawal from existing aquifers and discharges into drainage systems are closely regulated by state agencies. Water is stored and recycled to maximize usage. In some instances water is even returned to major aquifers through a series of recharge or connector wells. The phosphate industry has spent over 200 million dollars in the past decade for research, development and implementation of pollution abatement and water conservation systems.1 The following discussion outlines various phases of phosphate mining in central Florida today, with emphasis on research in waste disposal techniques and progressive land reclamation programs.

Jan 1, 1975

By Pierre Sauvé, Alex Makila

"The Camflo stock of porphyritic quartz syenite is an irregular pipe-like body 200 metres in diameter injected along the axial plane of an E-W trending F2 fold, near the contact between the. Kewagama and Malartic groups. The stock plunges 55° to the NE, almost parallel to the fold axis. The stock, exhibiting pervasive, weak cataclasis with numerous, locally auriferous quartz veinlets, is cut by minor faults striking WNW and dipping about 45°NE.Two types of gold ore are distinguished at Camflo: ""porphyry"" ore and ""fault zone"" ore. ""Porphyry"" ore, accounting for 90% of the gold produced, forms wide zones within the stock. The zones are irregular in shape but tend to be oriented E-W in plan view (on a horizontal plane) and are elongated paraliel to the plunge of the stock. These zones are more abundant in the northern part of the stock. Differences between ore and barren porphyry are not obvious. ""Fault zone"" ore occurs outside the stock along minor faults, particularly in differentiated sills offer rodiorite and felsic granophyre. Gold grade is roughly proportional to the pyrite content which in turns depends in part on the original magnetite content of the host rock. Although both types of ore show some correlation between Au and S, the Au/S ratio is considerably higher in porphyry ore.The bulk of the gold is in quartz syenite and in felsic granophyre, the most brittle lithologies at the mine. Presumably, the vertically continuous, pipe-like stock, made up of shattered rock, provided the main channelway for the ascending mineralizing solutions."

Jan 1, 1990

By Paul J. Smith, Eudomar A. Silva, Quang D. Tran, J. Milton Brooks

The Dallas Mill Creek Drainage Relief Tunnel is an 8.04 km (5 mile) tunnel that will provide 100-year flood protection for impacted neighborhoods in the east Dallas area. “Big Tex” is a main beam gripper TBM capable of changing diameters in mid-tunnel to excavate 2.83 km (9,290 ft) of 11.45 m (37.58 ft) diameter tunnel and 5.21 km (17,095 LF) of 9.93 m (32.58 ft) diameter tunnel. Construction began in April 2018 and is anticipated to be completed in 2024. This paper will discuss the challenges and lessons learned during assembly and re-assembly, start-up and mining with “Big Tex.”

Jun 13, 2021

By A. S. Choudhary

Most of the coal reserves in India are of drift origin and have very poor washability characteristics and a high level of near gravity materials. At least 30 new coal wash plants are expected to be set up in the next 5years.In such a scenario, adequate sampling followed by sink float test work for washability analysis becomes very critical for designing these new plants. For coal sampling purpose, stipulations of different standards are sometimes formidable and on the border of ?impracticality?; sometimes are self contradictory in terms of gross sample weight vis-a-vis minimum number of particles required in each size fraction and in each relative density fraction and also the minimum mass in the latter. Validation of the generated data is another important aspect of washability analysis. These issues can easily be addressed by using straight forward mathematical relationships, well known RR plots and by using simple MS Excel tool, etc.

Sep 1, 2012

By Katarina Theander, Robert J. Pugh

The generation of froth with a critical degree of stability and height is an important aspect of the froth flotation. In this paper we have studied the synergism and anti-synergism in frothing behaviour in a mixed flotation collector/frother mixture (a model fatty acid; sodium oleate and a model frother; nonionic (polyethyleneoxide alkyl C14E6). Values of the collector / frother molecular interaction parameters (the ß-parameters) were determined from the surface tension data using regular solution theory. The frothing characteristics of the system were evaluated using a glass column equipped with a series of conductivity electrodes, which enabled the water content of the froth to be determined. From these measurements, the dynamic frothability and static froth stability were determined. It was concluded that the synergism/anti-synergism in frothability was depended on the molecular composition of the surfactant system, the diffusion of monomer to the thin film interface and on the breakdown of the mixed nonionic/anionic micelles.

Jan 1, 2003

By M. Ashraf, P. J. Hazell, A. A. H. Ameri, J. P. Escobedo-Diaz, N. N. Elewa

A novel methodology to estimate the dislocation density in cubic metals by using microhardness measurements has been established. The proposed methodology is based on the Indention Size Effect (ISE) phenomena and strengthening mechanisms of materials at the micro-level. The methodology was validated by using experimental data of lean duplex stainless steel 2101 alloy. The results are confirmed via X-ray diffraction measurements and they substantiate that the proposed approach can be used as a general method to estimate dislocation density in metals.

Mar 1, 2017

Open pit mining at Bingham Canyon commenced in 1904 and about two billion tonnes of ore have been extracted since then. There remain just under one billion tonnes of ore in the pit and this will be mined by 2015. Because the porphyry system extends below the ultimate pit limits, Kennecott Utah Copper Corporation initiated a study in 1997 to determine the viability of underground mining. The study was initiated at that time because the results were to be used to compare the economics of a further pit expansion to that of an underground mine. If a pit expansion was viable then overburden removal would have to start in the near future. This paper is a summary of the work that was carried out for the underground study. Block cave mining was determined to be the most appropriate method to mine the deposit. Two block cave areas have been planned with a total production rate of 72 500 tpd. Additional sources of ore will come from skarn orebodies that are associated with the porphyry deposit. The geology of the deposits is described together with the geotechnical techniques that were used to characterise the rock in the absence of more detailed data that are normally available. A general description of the mine plan is given along with the reasons for selecting the particular methodology used. The study was completed at the end of 1998 and demonstrated the viability of underground mining at Bingham Canyon. Funds have been approved for further drilling and technical studies, which will lead up to a feasibility study in 2004/5.

Jan 1, 2000

By D. J. Boron, R. Kollrack

Introduction The US has the largest total (364 Gt, or 401 billion st) and recoverable (182 Gt, or 201 billion st) coal reserves in the world (Steam/Its Generation and Use, 1978). However, progressive coal use in this country has been hampered by environmental constraints, such as restrictions on sulfur dioxide, nitrogen oxides, and particulate emissions. Since the promulgation of the Environmental Protection Agency (EPA) New Source Performance Standards (NSPS), it has become apparent that all US coals will require sulfur removal to comply with the new sulfur dioxide emission levels. If more stringent emission standards are imposed, through pending acid rain legislation or amendments to the Clean Air Act, it is likely that the coal industry, and the industrial and utility sectors will look increasingly toward the development of new and advanced coal cleaning technologies to continue and promote the use of coal. Table 1 summarizes the EPA NSPS for sulfur dioxide emissions. Fortunately, the basis for determining sulfur dioxide removal is the Btu and total sulfur content of the run-of-mine coal. Thus, credit is given for any sulfur removed before combustion, e.g., through physical coal cleaning. Unfortunately, physical coal cleaning alone may be unable to produce a compliance coal with respect to the EPA NSPS. Therefore, all industrial and utility coal-fired boilers will require post combustion gas cleanup, e.g., flue gas desulfurization (FGD), to comply with the new regulations. Table 1 - EPA NSPS for SO2 Emission Category SO2 Content of ROM Coal SO2 Reduction Required 1 0-2 Ib/M Btu 70 2 2-6 Ib/M Btu To 0.6 Ib/M Btu (70% to 90%) 3 6-12 Ib/M Btu 90% 4 12 Ib/M Btu To 1.2 Ib/M Btu Alternatives to post combustion gas cleanup, such as staged combustion, lime and coal injection, fluidized-bed combustion, and chemical coal cleaning, are currently under development. Of these, chemical coal cleaning offers the advantages of providing (1) high quality fuel for combustion (low sulfur and ash, high Btu, and attenuated product variability); (2) a solid fuel product for use in retrofitted gas and oil boilers; and (3) a solid fuel product with multiple end uses, e.g., internal combustion engines, carbon electrodes, coke, or chemical feedstock. Gravimelt process In the Gravimelt process, 1.4 mm x 0 (14 mesh x 0), physically beneficiated coal of about 5% moisture is fed into a 3401 to 3901 C (644 ° to 734 ° F) molten bath of sodium hydroxide and potassium hydroxide. Organic and pyritic sulfur and most of the mineral matter in coal react with and are dissolved in a molten mixture of potassium hydroxide and sodium hydroxide. The reacted coal sulfur is converted to alkali sulfides and polysulfides, whereas the mineral matter is apparently converted to alkali-aluminum silicates and other byproducts that are water and acid soluble. Coal residence time is 60 to 240 minutes. The sodium hydroxide to potassium hydroxide ratio is about 9:1. The total hydroxide-to-coal ratio is no less than 10:1. The specific gravity of the caustic melt is about 1.8 g/cm3 . Thus, coal is buoyant in the bath, and insoluble mineral matter with density greater than 1.8 g/cm3 would sink. Reacted coal is first water-washed with an equal weight volume of water in a countercurrent mode of operation. The coal is filtered and the filtrate sent for regeneration. The coal is then acid-washed in 10% sulfuric acid. The acid-water/coal mixture is filtered to recover the final coal product, and the filtrate treated with lime to precipitate dissolved mineral matter. The precipitate is filtered from the mixture and discarded. The water is recycled to the plant. Filtrate from water-washing is treated to remove sodium sulfide. The sodium sulfide is chemically treated to produce hydrogen sulfide, calcium car¬bonate, sodium hydroxide, and other products. The hydrogen sulfide is treated to produce sulfuric acid for the acid-washing step. The calcium carbonate is treated further or discarded, and the sodium hydroxide is concentrated by evaporation and recycled to the reactor. The Gravimelt process is currently at the bench-scale stage of development. A 0.9-kg (2-lb) continuous reactor is in operation with a continuous water- and acid-washing section adjacent. The regeneration section is not developed to a meaningful extent, although key reactions have been tested in the laboratory. Process conditions, product recovery efficiencies, etc., however, have not been determined adequately and continuous regeneration operation has not been attempted. Work is now in progress to develop each section to a continuous bench-scale operation in processing coal at 9.1 kg/h (20 lb per hr). Microwave process The carbon and hydrogen in coal are relatively transparent to microwave radiation. However, water, caustic, pyrite, and other mineral matter components are strong conductors of microwave radiation energy at the proper frequency. As a result, selective heating of the noncarbonaceous materials occurs that enables the chemical reaction of sulfur and ash with caustic. Although very high temperatures may result from

Jan 1, 1987

By Ham. Y. M.

The corrosion behaviour of Al-B4C metal matrix composites immersed in 2500 ppm boron-containing H3B03, 0.5 M K2SO4, and 3.5% NaC1 solutions was investigated. Effects of B4C particle volume fraction on the corrosion behaviour of composites were studied using potentiodynamic polarization and electrochemical impedance spectroscopy techniques. Scanning electron microscopy was employed to study the surface morphology of the material before and after corrosion. Results show that all materials investigated are more corrosion resistant in 2500 ppm boron-containing H3B03 than in other solutions. No appreciable corrosion was observed in boric acid and sulfate solutions in contrast to an obvious pitting in NaC1 solution for both base alloy and composites. The preferential pitting sites are the interfaces of Al/Fe intermetallics for the base alloy, and the A1/B4C interfaces for the composites. Increasing the volume fraction of B4C particles in the composite decreases its corrosion resistance of Al-B4C composites.

Jan 1, 2011

By Herry Poernomo, Agus Prasetya, Iga Trisnawati, Himawan Tri Bayu Murti Petrus, Gyan Prameswara, Panut Mulyono

We studied yttrium extraction from untreated zircon sand processing waste tailings. Xenotime mineral as the rare earth element source with an abundance of yttrium (REY) was detected in sufficient grade for extraction. As much as 9.03% yttrium exists in zircon tailings (analysed by X-ray fluorescence spectrometry). The presence of yttrium was confirmed by X-ray diffractometry in xenotime minerals, which are yttrium carriers (Y-PO4). The purpose of this research was to determine the effect of leaching conditions on yttrium recovery from zircon sand after alkaline fusion treatment. Alkaline fusion was chosen to decompose phosphate into hydroxide in the xenotimemineral,which will reduce further required hydrometallurgical processing.Alkaline fusionwas carried out for 3 h at 450 °C, with a ratio of sodium hydroxide to zircon tailings sand of ~ 1:1. The alkaline fusion product was leached with water, followed by hydrochloric acid treatment to leach the yttrium. Yttrium recovery reached 87% under optimum conditions (60 °C, 1MHCl, and solid-to-liquid ratio = 1/10 for 7.5 min). A suitable model for yttrium dissolution with hydrochloric acid was diffusion through a solid particle ash layer. The calculated activation energy (EA) for this model was 20.21 kJ/mol.

May 26, 2020