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By Jewel A. G. Gomes

Electrocoagulation (EC) has been performed with Al-Fe electrode system with changing polarity for the removal of arsenic from water. The removal process has been studied using ICP-AES and Anodic Stripping Voltammetry with a wide range of arsenic concentration (1-1000 ppm) at various pH (4-10) and residence times (1-60 min). Electrochemically generated by-products have been analyzed using XRD, FTIR, XPS, SEM/EDS and Mössbauer Spectroscopy. Results revealed expected crystalline iron oxides (magnetite (Fe3O4), lepidocrocite (FeO(OH)), iron oxide (FeO)) and aluminum oxides (bayerite (Al(OH)3), diaspore (AlO(OH)), mansfieldite (AlAsO4?2(H2O)), as well as some interaction between the phases. The amorphous or ultra-fine particular phase was also found in the floc. The substitution of Fe3+ ions by Al3+ ions in the solid surface has been observed, indicating an alternative removal mechanism of arsenic in these metal hydroxides and oxyhydroxides by providing larger surface area for arsenic adsorption via retarding the crystalline formation of iron oxides.

Jan 1, 2006

By O. A. Bascur

The problem of setting the proper control objectives and/or having several objectives to be met in the implementation of grinding control strategies can be resolved. The design and implementation of a distributed control system for safe operation of a grinding circuit while maximizing profit is presented. An optimal recovery/throughput target for a desired grind is accomplished by advanced control strategies at the DCS level. A mill load constraint control technique provides stable operation with a substantial increase in throughput, while the desired grind index is maintained by an inferential grind control technique without using a particle size analyzer. The integration of crushing/grinding and flotation in the overall design was the key to the success of the project. This synergistic effect was obtained by implementing a computer -plant management system in the concentrator. The overall philosophy of keeping the implementation simple but not simpler was used.

Jan 1, 1989

By W. M. Albrecht, W. D. Klopp, R. I. Jaffee, B. G. Koehl

Kinetic studies were made of the reactions of tantalum with oxygen, nitrogen, and air at 400o to 1500°C. The tantalum-oxygen reaction is linear from 500° to 1250°C. The tantalum-nitrogen reaction follows a cubic rate law at 400° to 700°C and a parabolic rate law at 800" to 1475°C. The reaction behavior in air is initially parabolic followed by a transition to linear behavior. In the search for high-temperature materials, considerable interest is being shown in the refractory metal, tantalum. It is the third highest melting metal with a melting point of 2996°C. Also, tantalum has excellent ductility, making it one of the most workable refractory metals. In considering high-temperature applications of tantalum, the oxidation kinetics of the metal must be known. Such information also serves as a basis for the development of corrosion resistant tantalum-base alloys. Therefore, this study was made to investigate the reaction of tantalum with air, nitrogen, and oxygen. LITERATURE Several oxides of tantalum have been reported. The pentoxide, Taz05, is the only oxide of tantalum whose identity has been established. A high-temperature modification, aTa2Os, has been reported by several investigators.1-4 A low-temperature phase, ßTa2O5, undergoes an allotropic transformation at 1320o to 1350oC. 5-8 Four other oxides of tantalum have been reported. The oxides TaO2, TaO, and Ta4O have been reported by schonberg.9 A suboxide Ta2O, was observed by Wasilewski 6. However, the presence of oxides lower than Ta2O5 is questioned by Lagergren and Magneli.7 In oxidation studies, vermilyeal0 and Peterson, et al.,12 reported one possible phase intermediate between Ta and Ta2O5. Only Ta2O5 was formed in the tantalum air reaction, as reported by Michae1.l3 Several studies have been made on the kinetics of the tantalum-oxygen reaction. vermilyeal4 investigated the range 50° to 300°C and found that the rate of growth of a thin adherent oxide film could be described by the Cabrera-Mott theory. Gulbransen and Andrew15,18 studied the reaction kinetics of tantalum in oxygen pressure of 1/10 atm at 250° to 450°C. At all temperatures the reaction followed the parabolic rate law for periods up to two hours. The oxidation behavior in the range 500" to 1000°C was found to be linear at oxygen pressures of 0.2 to 600 psi according to Peterson. et al.l1 At very low pressures (1 x 10-3 to 2 X lo-' mm of Hg) Gebhardt and seghezzi17 reported linear oxidation at 800° to 1500°C. The reaction of tantalum with air is similar to the reaction with oxygen. Studies have been made by Waber, et a1.,12 and Michae1.l3 Two nitrides of tantalum, TaN and Ta2N are known. Their structures have been determined by Brauer and Zapp 18,19 The tantalum-nitrogen reaction kinetics have been investigated by Gulbransen and Andrews. 15, 16 The reaction follows the parabolic rate law at 500° to 850°C at a nitrogen pressure of 1/10 ,atm. At the same temperature, the tantalum-nitrogen reaction is much slower than the tantalum-oxygen reaction. EXPERIMENTAL TECHNIQUES Material— The material used in this study was high-purity, electron-beam-melted tantalum obtained from the Temescal Metallurgical Corp. The hardness of the as received ingot was 76 Vhn. The analysis of the material is given in Table I. Methods—Cylindrical specimens of tantalum approximately 0.9 cm in diam and 0.5 cm in length

Jan 1, 1962

By C. P. McCormack

THE Cuyuna manganiferous iron ores can be a principal source of manganese for the iron and steel industry in the United States, provided metallurgical methods as a whole are adjusted so as to use run-of-mine ore. The use of manganese in steelmaking is growing. Increasing amounts of impurities in raw materials call for more manganese, a constant cheap supply of which is wanted. There are few deposits of ore in the United States carrying over 35 per cent. manganese that can be mined and marketed at a profit in normal times. We are largely dependent on foreign ores. The small domestic deposits should rightly be held in reserve for an emergency, such as a war, when importation might cease. Even the production from lean ores, concentrated it may be, might not carry us through a protracted stringency. The Cuyuna district is, therefore, during normal times capable of supplying a large total of manganese. The ores are iron ores containing 5 to 25 per cent. of manganese with the inverse range of 52 to 28 per cent of iron. The manganese costs no more per unit in the ore than the iron, because cost of mining, largely by open-pit methods, is not affected by the manganese content. The manganese units should, perhaps, command a higher price, but as their value is contingent on metallurgical practice, a higher price may be justified only when a shortage in manganese occurs. Manganiferous iron ores can never entirely supplant straight manganese ores though methods are employed that will permit them to do so partly. For the manufacture of ferromanganese, ores carrying above 42 per cent. manganese are needed. Ores with manganese running between 20 and 42 per cent. are used for making spiegeleisen. But established practice in certain plants, based on the admixture of manganiferous iron ores to straight iron ores has greatly lessened the consumption of high-priced, scarce, manganese alloys with good results.

Jan 2, 1925

By Mark A. White

Arch Mineral has developed several computer programs that aid the surface mine planning function. These programs are utilized to perform long term projections for various area or semi-area stripping operations. The program generated reports include yardage, tonnage, quality and ratio information for the user-specified time periods. The projections are used to anticipate the annual budget, to compare alternative mine plans and to analyze acquisition candidates. The programs are user friendly, fast and accurate. Geological Grid File Two data files must be set-up and used in conjunction with the mine planning programs. These are a geological data file and a general input file.

Jan 1, 1984

By I. Ohnaka

Although many works have been reported on the semisolid casting, the mold filling behavior of semi-solid alloys has not been so well understood. This paper compares the physical and numerical simulation results on the mold filling behavior. Physical simulation was performed by using solid-water mixture. The numerical simulation was made under the assumption of one phase flow with an effective viscosity. It was found that the solid particles stabilize the flow greatly if the kaction solid exceeds about 0.4, because viscosity increases greatly. The overall flow behavior can be estimated by the numerical simulation. Interesting flow behavior was also found which can not be observed in the case of one phase flow and which is difficult to simulate by the one phase flow assumption.

Jan 1, 1997

Flotation of Phosphate Gangue from Magnetite Fines - Non-Ionic Surfactant as Atrac Collector Modifier

Sep 13, 2010

By Alfred Weiss

Selecting a minicomputer is not unlike selecting a costly piece of mining equipment such as a truck, shovel, crusher, or conveyor belt system. A marked difference might lie in the frequency at which such selection takes place; another difference might be mine management's familiarity with mining equipment vis a vis a minicomputer. These basic differences might prompt a mining company to seek the aid of a consultant when selecting a minicomputer system. Advantages and Disadvantages The fundamental advantage of employing an independent consultant is that it provides management with a source of knowledge and breadth of experience not readily available within the organization. If the consultant has a minerals background, such knowledge and experience are further enhanced by actual experience in exploration, mining, and mineral processing. If the consultant has worked within a mining company, such hands-on experience is supplemented with a meaningful understanding of the client's business; this facilitates communication between the consultant and members of the client organization. The consultant has been, so to speak, "in the client's shoes." There are several other advantages. Since computer systems cut across organizational boundaries, the independent consultant may be able to move within the organization more easily than an inside staff person to collect necessary information. Also, detachment from any manufacturer should provide the client with an objectivity and a set of recommendations as near optimal as feasible. The principal disadvantages lie in the area of continuity and accountability. Once the consultant's report has been submitted and his recommendations implemented, a discontinuity might occur. Also, if implementation does not provide the desired results, the consultant might not be available to correct the situation or account for actions taken. Proper selection of consultant services, relative to kind and duration, and the consultant's method of operation, will often prevent many of these aftereffects. Services Offered The professional consultant can help a client: •Define and summarize information systems technology* needs; • Evaluate alternatives such as in-house capability and capacity, time-sharing, or service bureaus; • Prepare a formal request for proposal (RFP) to be submitted to vendors; •Evaluate vendor proposals; •Evaluate the economics of lease vs. purchase; and •Assist in the conversion and installation. Thus, consulting services may range from evaluating an existing system and upgrading to a larger piece of equipment, to selecting a new first-for-the-organization computer. First, Define and Summarize Needs Regardless of circumstances, the most valuable document required before equipment selection is the information system plan (I/S plan). It helps insure that the company's information needs are met in a cost-effective manner. An I/S plan is usually a comporient of the mining company's overall business goals. Since an I/S plan is a formal vehicle for establishing and communicating the direction the organization intends to take to satisfy information requirements needed to achieve goals and objectives, most initial systems services are likely to center around creating such a document or updating an existing document. The trend toward company-wide I/S planning is increasing in the mining industry; however, smaller scale efforts focused at the information systems needs to support an individual

Jan 11, 1981

By S Farrokhpay, C Akop

Copper-gold-pyrite ore types with an elevated pyrite head grade utilise a sequential flotation circuit to recover initially a copper-gold concentrate with a pyrite-gold concentrate being produced from the tailing of the initial circuit.A simpler flotation circuit with lower operating and capital costs for treatment of such copper-gold-pyrite ores with an elevated pyrite head assay was sought in the described work. After characterising the ore, the objective was to apply bulk rougher flotation of both the copper minerals and pyrite for the copper-gold-pyrite ore at natural pH. Hence, no lime or other pyrite depressants were required in bulk roughing. The bulk rougher concentrate was upgraded in a cleaner circuit to reject non-sulfide gangue and then the resulting sulfide concentrate was separated by depressing the pyrite to produce two products (copper and pyrite concentrates, both having significant gold assays). A standard test with a sequential circuit was applied to the same ore for comparison.The chalcopyrite and pyrite losses in the tailing from the rougher circuit were the same for both bulk and sequential roughing (less than six per cent). In cleaning (bulk circuit), the effective pH for pyrite depression was 11.8–11.9; copper recovery was 70–75 per cent (grade of 15–20 per cent copper) and the sulfur recovery (pyrite product) was 80–85 per cent (grade of 34–36 per cent sulfur).These results approached those for the sequential circuit where copper recovery was 80–85 per cent (grade of 15–20 per cent copper) and the sulfur recovery (pyrite product) was 83 per cent (grade of 38 per cent sulfur).The observed effects of feed sizing on the liberation of minerals in the rougher feed are discussed. Options for improving the bulk circuit results are also considered. Importantly, the implications of various aspects of these findings for the bulk circuit are discussed with an industrial focus.CITATION:Akop, C, Farrokhpay, S and Johnson, N W, 2014. An alternative circuit for chalcopyrite-pyrite ores with elevated pyrite content in Cu-Au ore treatment, in Proceedings 12th AusIMM Mill Operators’ Conference 2014 , pp 407–416 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 1, 2014

By B Madeja-Struminska

The method of evaluation of the approximate temperature of arising spontaneous fire centre and time of coal self-ignition is presented in this paper. The method uses BystronÆs (1997) model of coal self-heating and measurement of carbon monoxide concentration in the mine air outside the fire centre. Knowledge of the spontaneous fire centre temperature and time of fire beginning allows for more accurate analysis of the fire hazard in the coal mine and allows for adequate time to undertake special fire prevention actions. The method is illustrated by an example from Polish mining practice.

Jan 1, 2005

By Joel N. Van Sant

An investigation was conducted to study possible sources of Utah refractory and structural clays that could be used for the manufacture of refractories and light or buff burning heavy-clay products. A field study was made and the clay samples tested to determine their firing characteristics. Results of the Bureau's work indicate many small noncommercial deposits of hypogenic clay occur in igneous rocks. Commercial hypogenic deposits are mostly halloysite clay developed by hydrothermal replacement of limestone. The high firing shrinkage of halloysite precludes its use as a refractory, however, halloysite produced at the Dragon Mine south of Eureka, is used for making oil catalysts. Clays and shales suitable for low heat duty refractory products are found in the Manning Canyon shale, in shales associated with coal bearing rocks, and in the Dakota formation. Some of the clays from the Dakota formation are also suitable for high-duty refractory products.

Jan 1, 1964

By J. H. Viljoen, A. Heyns, J. F. Morrison, C. H. Wyndham, N. B. Strydom

Discussion M. J. Martinson* (Associate Member): This is another interesting paper in the series which the Human Sciences Laboratory is publishing on the physiological cost of the more strenuous manual tasks performed underground. As a mining engineer-albeit presently an academic one -this contributor is more concerned with the practical significance of the study in the field of stope cleaning than with physiological niceties, but before discussing some of the practical issues raised by the paper perhaps the authors would care to reply to lay comment on three physiological matters. Firstly, in two places the authors make statements to the effect that the maximum oxygen intake of 'average' (sic) Bantu mine workers is 3.0 l/min. What do the authors mean by 'average'? Maximum oxygen intake in the individual is probably some function of inter alia age, physical fitness, state of health, ambient partial pressure of oxygen, diurnal and/or seasonal changes in metabolism, and possibly cumulative exposure to dust and other occupational contaminants. Since maximum oxygen intake is apparently such an important quality in manual labourers we ought to have a clear picture of its distribution in the entire labour force, and also know precisely how the individual's quantum is affected by factors such as those mentioned above. Secondly, arising out of the same statements and ignoring the reservation implicit in the previous paragraph, is it correct to say that the 'average' Bantu mine worker has a maximum intake of 3.0 l/min? In the August 1966 issue of the South African Journal of Science four of the present co-authors reported the mean maximum oxygen intake of a random sample of 338 Bantu recruits of different tribes to be 2.63 l/min with a standard deviation of 0.542 l/min, and these figures were repeated in a paper published in the November 1968 issue of this Journal. Admittedly the mean maximum oxygen intake of the six subjects used in the present study was 3.05 l/min, but since they were selected on the grounds of previous shovelling experience they can hardly be regarded as being 'average' recruits. Thirdly, it seems to me that the authors are perhaps a little uncritical in their acceptance of the '50 per cent of maximum oxygen intake' level of activity for endurance work. I am not unmindful of the evidence upon which this level is based, but I suggest that the evidence is at best tenuous and that the whole topic of fatigue and 'optimum' levels of productivity under conditions of high muscular activity and considerable psychological stress is a lot more complicated than would appear from the authors' ready acceptance of the 50 per cent level. Under present working conditions the question is largely rhetorical because it would need an army of supervisors to ensure that all members of a typical gang work at 'optimum' levels throughout the shift, but if individual performance should ever become more readily determinable the issue might become crucial to management and malayisha alike. So much for physiological matters; now a few random thoughts on some wider, practical issues raised by the paper. In the paragraph headed Material the statement is made that 'The mixture (of fine, medium-sized and bigger rocks) was considered to be representative of an average underground rock sample'. Elsewhere in the report the authors specifically draw attention to the effect of particle size on the oxygen consumption of men shovelling graded material, and intuitively one imagines that the size distribution of the material being shovelled would be highly significant in a study such as this. Presumably the authors were forced to assess size distribution subjectively because of the lack of objective data on the subject, and possibly the necessity for this assumption highlights how little we know quantitatively about any of the variables in stope cleaning systems. In this connection it may be remarked that although the present study was largely designed to show how dip and stoping width effect energy expenditure in shovelling, no data are given-nor, probably, are they available-to show how the industry's stoping tonnage is distributed in terms of these two factors. It seems to me that if the mining industry seriously hopes to improve stope productivity it must start by measuring the basic variables in existing systems. Elaborating somewhat on the previous paragraph, there also seems to be a tremendous gap between the present study-and its predecessors-and the prospect of gainfully using the information so carefully garnered by the authors. It would be reassuring to hear from the authors that their work on the physiological aspects is being integrated in some master plan for increasing productivity of manual labour. This I fear is not the case, and if productivity has changed at all in recent years I would hazard a guess that the change has been brought about by random factors rather than by systematic engineering design. The Human Sciences Laboratory has worked extensively in stopes in the past, and possibly the time has come when it should widen the ambit of its operations to include engineering-type analyses of cleaning systems, with particular emphasis on the economics of rapid cleaning by manual labour. Perhaps experiments could be conducted with stakhanovite-type gangs consisting of men selected for their high maximum oxygen intake and superior motivation; these men might be offered special incentives in the shape of extra pay

Jan 10, 1969

By Theo. Koulomzine

"The property of Consolidated Central Cadillac Mines, Limited, consists of 1,103 acres formerly held by Central Cadillac Mines, Limited, and Wood Cadillac Mines, Limited. Both mines have been partly developed to depths of 565 and 1,000 feet respectively, and from 1939 to 1943 they produced gold, with small amounts of silver and tungsten, having a total value of $2,208,528 from 377,489 tons of ore. The average grade of the ore was $5.85 a ton . Less than 10 per cent of the ore hoisted was discarded by sorting. New development work, carried out since the former Wood property has been purchased by Central Cadillac, has shown that the tonnage indicated by mine workings and diamond drilling exceeds 2,000 tons per vertical foot.The property lies between the claims of Kewagama Gold Mines, Limited, and Pandora Cadillac Gold Mines, Limited, and straddles the Cadillac 'break' for a length of nearly 1 ½ miles. The main workings are 1 3/4 miles east of the well-known O'Brien gold mine.GEOLOGYOutcrops are very scarce, with none at all of the most important horizons, so that the general geology shown in Fig. 1 is deduced mostly from underground workings and diamond drilling. Fig. 2 is a vertical cross-section and Fig. 3 a longitudinal section through the ore zone.The ore lies within a belt of Blake River volcanics which is approximately 600 feet wide and is flanked on the north and south, respectively, by Cadillac and Kewagama sediments. All the formations have a nearly vertical attitude. The Blake River belt consists essentially of well sheared volcanic tulfs, but associated with these in the northern part are a number of bands of iron formation and the central part, to a width of 100 feet, is occupied by talc schists, making the Cadillac fault proper. Intruding the volcanics south of the Cadillac shear zone are dykes of two types- finegrained aplite or quartz albitite, and porphyry; both consist essentially of albite and quartz and are distinguishable only by the grain size. They have the general east-west trend of the formations they intrude."

Jan 1, 1954

By H. M. Chance

By courtesy of the officials of the Chicago, Burlington and Quincy railroad, I am permitted to publish the results of an examination made in 1887 and 1888 of' the country west of the Black Hills, and lying between the Black Hills and the foot-hills of the Big Horn Mountains in the northeastern part of Wyoming. With the exception of the country constituting the foot-hills, or the encircling ridge or rim of the Black Hills proper, all of this area is occupied by rocks of Cretaceous age. In the central part of the Black Hills are found the tin-bearing granites, the mica-schists, and gold-bearing slates, already so well known. Surrounding this central area of older rocks (which should be considered as the "Hills" proper) are the attenuated representatives of the Palzeozoic, dipping gently away from the Hills in every direction, overlaid by the gypsum-bearing Mesozoic, which here forms a bold high ridge or rim, completely encircling the Black Hills, and known as the inner ridge. Overlying, and apparently conformable to this, we find the Dakota sandstone, which likewise forms a prominent ridge or rim, known as the first or outer ridge. This is the formation which contains the coal now worked at the mines near Newcastle. It is also the formation in which the best oil-wells have been obtained in the "Stoakdale," or "Beaver," and the " Belle Fourche " oil-districts. Overlying it are the Colorado shales, forming the great plains drained by Beaver creek and the south fork of the Cheyenne river, and extending west to the Belle Fourche, near Donkey creek, where the overlying Fox Hill formation (thin) and the Laramie group, the great lignite-formation, are first seen. The Laramie group occupies the surface from the Belle Fourche at Donkey creek west to the foothills of the Big Horn mountains at Buffalo, an air-line distance of some seventy or eighty miles.

Jan 1, 1891

By Priscila Tavares de Paula, Márcia Mika Saito, Henrique Gonçalves Teixeira, Felicio de Freitas Simoões

In this paper a sample of the apatite flotation from Cajati-SP was studied. Four alternatives of depressants were analyzed: corn starch, cassava starch, sodium hexametaphosphate with corn starch and sodium silicate with corn starch. The cassava starch had a gain of 7.5% in the metallurgical recovery and a collector consumption 13.8% greater than corn starch. Alternatives with hexametaphosphate and silicate sodium had a loss of 6.7% and 7.8% in recovery respectively in comparison with corn starch.

Oct 3, 2019

By I Hustwick

New Zealand Steel Mining Limited operates two independent sand mining operations on the West Coast of the North Island of New Zealand. Waikato North Head 30 km south of Auckland supplies ironsand concentrate as the raw feed for New Zealand Steel's works at Glenbrook and Taharoa 70 km further south, exports ironsand concentrate to Japan.   Taharoa uses two sets of six low pressure Warman centrifugal pumps in series to load bulk ore carriers 3 km off shore through two underwater pipelines. Waikato North Head delivers the ironsand to Glenbrook 18 km through a 200 mm ID buried pipeline using two stages of high pressure positive displacement (HPPD) pumping.   A high level of operating and maintenance expertise has been developed since commissioning the two systems which are proving to be reliable, efficient and of minimal impact environmentally.

Jan 1, 1990

By John Marsden, Mike Botz

A variety of modeling techniques can be utilized to forecast metal production at heap leaching operations. These approaches reflect a wide range of complexity, flexibility, time to implement, cost, and accuracy. For many operators, a spreadsheet-based modeling technique is attractive since the calculations are directly accessible, models can often be developed by site staff, and the results are generally easy to extract and interpret. The authors have developed a spreadsheet-based modeling technique that provides a high degree of flexibility, while still considering detailed operating information for ore properties, leach kinetics, scaleup factors, lift height, and in-heap metal inventories. The technique involves establishing kinetic leach curves for each ore type to characterize metal extractions, followed by application of separate metal recovery curves to define the rate at which leached metals exit the heap in pregnant solution, taking into account delays due to the in-heap solution inventory. At any point in time, the difference between total metal extracted and total metal recovered is equal to the in-heap inventory of the leached metal. This modeling technique is particularly useful for larger multi-lift heaps where delays in metal recovery are appreciable due to in-heap solution holdup. The technique is also applicable to sites with multiple heaps, leach cycles, and/or metal recovery plants (e.g., carbon columns for gold; solvent extraction for copper). This paper describes the modeling technique, including input data required and methods for defining kinetic leach curves and metal recovery curves. The technique is compared against an alternative spreadsheet technique using a simplified in-heap solution inventory calculation to forecast metal production, which has proven to be effective at a number of sites.

Sep 3, 2019

By G. L. F. Powell

g samples of Cu-20 wt pct Ag alloy have been mdercooled to a maximum of 197°C by melting under a slag of commercial soda-lime glass in a vitreous silica crucible. No grain refinement of the primary copper was observed in samples undercooled to the maximum of 197°C. When the samples contained a small amount of oxygen, the copper dendrites were partially recrystallized at undercoolings greater than 97°C. In previous papers'-3 reporting the grain structure of undercooled silver and copper, it was observed that grain refinement was dependent on both undercooling and oxygen content. Grain refinement occurred in undercooled silver when the degree of undercooling exceeded the range 153" to 175"C, while in Ag-0 alloys (0.12 wt pct) fine equiaxed grains were exhibited when undercooling was greater than 50°C. Similarly, copper samples undercooled as much as 208°C displayed fan-shaped growth from a single nucleation site, while the grain structure of Cu-O alloys (0.08 wt pct) was fine and equiaxed at undercoolings larger than 150°C. Thus the presence of oxygen greatly reduced the undercooling at which grain refinement occurred. It was also observed that the change in grain size resulted from recrystallization and was not due to an enhanced nucleation rate in the liquid-solid transformation. It is possible that the influence of oxygen on recrystallization is due primarily to its presence as a solute element. walker4,' reported that, although a grain size change did not occur in pure nickel until the undercooling exceeded 150°C, small grains were observed in samples of Ni-Cu alloy solidified at small and large degrees of undercooling. Jackson et al.6 suggested that the fine grained structure of the Ni-Cu alloy resulted from the melting off of dendrite arms during recalescence. This remelting process may occur in alloys as a result of segregation during freezing which causes a variation in liquidus temperature from point to point within a dendrite. It was therefore decided to undercool copper with a metallic alloying element to ascertain whether the presence of a metallic solute would have a similar effect to oxygen in inducing grain refinement. A Cu-Ag alloy was chosen, since both metals had been shown to behave similarly on undercooling. The alloy Cu-20 wt pct Ag was selected since the eutectic constituent outlines the initial growth form of the primary copper, so that the as-frozen grain structure is not obscured if subsequent recrystallization occurs. This paper describes the results of undercooling experiments carried out with Cu-20 pct Ag samples undercooled to a maximum of 197°C and the effect of oxygen content on the grain structure of the undercooled samples. EXPERIMENTAL Melting was carried out in a small cylindrical resistance furnace using "fine" silver granulate and oxygen-free high conductivity copper. The procedure adopted was to melt the required quantity of silver in air in a clean vitreous silica crucible for approximately 15 min, freeze, and add granulated commercial soda-lime glass to form a complete surface slag cover, after which the sample was melted and frozen several times to reduce the oxygen content. The glass slag cover was approximately 3 in. thick. Pieces of copper (=50 g) were added to the crucible until the required quantity to make 350-g samples of alloy had been charged. Each piece was added quickly to the crucible which was held at a temperature slightly above the melting point of silver. The piece was quickly pushed beneath the glass to minimize oxidation and any oxide coating usually decomposed before the piece had settled down into the silver. After the full quantity of copper had been added, the melt was stirred with a silica rod to hasten homogenization and a Pt/Pt 13 pct Rh thermocouple enclosed in a vitreous silica sheath inserted for temperature measurement. Heating and cooling curves were recorded on a potentiometric chart recorder fitted with a zero suppression unit. The milli-voltage range of the recorder was adjusted so that temperatures could be read to 1°C. Heating and cooling curves were taken every hour until three consecutive readings gave the same solidus-liquidus range, consistent with the solidus-liquidus range for this alloy composition by reference to Hansen and Anderko.7 Metallographic examination of samples frozen at this stage, failed to show any variation in composition from bottom to top of the ingot. Consequently, it was considered that the melt was homogeneous at this stage and undercooling experiments were then car-

Jan 1, 1970

By R Redden

"The Didipio Gold-Copper project in the Philippines has been a well-defined reasonably high gold grade porphyry for almost two decades. OceanaGold are now bringing the project into production in late 2012 and have recently reoptimised the project with the goal of combining the inherent high grades with higher through-put than previously proposed that will see strong initial metal production. The project was discovered and well defined by 1995, but has seen some obstacles to development, namely sovereign risk, capital cost and project funding. The project strengths over this time that have kept the flames alive are: strong grades large orebody with low stripping ratios a simple beneficiation process and a clean high-grade concentrate strong support from the Philippines government.The initial study in 1998 used Gold and Copper price assumptions of $325/oz and $0.95 lb respectively. Today reserves are based on $950 and $2.85. So, being a porphyry with a large reducing grade halo, the deposit now lends itself to extraction on a much larger scale than before, without any significant capital expenditure as a larger open pit now dominates the project.Didipio is a game changer for OceanaGold, so it has been important to learn from history and to capitalise on a strong initial metal production profile. The project sits in a reasonably remote mountainous area of Luzon which sees annual typhoons and numerous short duration high intensity rainfall events and two rivers flowing through the project area, so it has not been without some innovative technical solutions.CITATION:Redden, R, 2012. Didipio - Learning the lessons, adapting and getting ready for production, in Proceedings Project Evaluation 2012, pp 67-76 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

May 24, 2012

Economic Sustainable Power for Mineral Processing

Sep 13, 2010