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By C. G. Carey, M. P. Mullins

BHPBilliton is a large (US$32bn market capitalization) diversified mining group. Significant organic growth is planned through the development of a number of mining, processing and oil and gas projects, totalling some US$5.2bn over the next three years. The investment standards underpin the capital approval process, and incorporate the requirements for range and risk analysis of key input variables, including mineral resources. Public reporting codes for mineral resources and ore reserves (such as JORC) stipulate deterministic reporting of estimates, but are generally silent on any requirement for range or risk assessment, although they acknowledge the risk and uncertainty in the production of resource estimates. In contrast, public reporting of petroleum reserves allow for the use of either deterministic or probabilistic estimation, and for the latter, mandate the level of confidence required for proven versus probable reserves. Conditional simulation techniques, popular in the petroleum industry, provide an ideal tool forrange and risk assessment of the range of outcomes of key resource parameters, such as tonnagesand grades. Examples of a diamond deposit and of a porphyry copper deposit are presented. The understanding and quantification of mineral resource risk through a technique such as conditional simulation provides quantified information on the potential upside or downside. These techniques should be more widely utilized in the high-risk minerals extraction industry. Keywords: mineral resources, conditional simulation, risk, capital investment standards, BHPBilliton.

Jan 1, 2003

By A M. Garforth

Insurers are hesitant to provide coverage for Open cast Pit side wall collapse, and even the consequent damage to plant and equipment in the pit as well as the resulting interference with the mining operations. After a general introduction to relevant principles of insurance, this paper addresses some of the major reasons for that reluctance based upon the authors? extensive claims experience, examines what limited coverage can still be obtained and looks to the future possibilities. Confidentiality of sources and examples referred to in this presentation is a natural imperative so that a bibliography is not provided but reference is made to relevant sample insurance policy wordings.

Jan 1, 2006

By Zay Jeffries

THIS paper will include a brief discussion of Prof. Howe's paper on The Supposed Reversal of Inheritance of Ferrite Grain Size from that of Austenite.1 The general subject of grain refining in steel and iron will also be treated. The conclusions reached are stated in the following hypotheses which are discussed more in detail below. 1. The ferrite grain size in pure iron, the ferrite and pearlite grain size in hypoeutectoid steel, the pearlite grain size in eutectoid steel and the cementite and pearlite grain size of hypereutectoid steel are not inherited from the grain size of the mother austenite. 2. The only structural feature that is generally inherited from the austenite of hypo- and hypereutectoid steels on cooling through their transformation ranges is the position of the excess ferrite or cementite at the austenite grain boundaries, sometimes causing complete and sometimes incomplete networks which outline the old austenite grain boundaries. Rapid cooling through the transformation range will cause the non-inheritance of this structural (network) feature. 3. The austenite grain boundaries themselves are nearly always effaced in all steels and also in pure iron during the Ar transformations. 4. The grain-size refining of steel and iron is brought about by the combined. effect of non-inheritance of the transformation products on either heating or cooling, i.e., the austenite transformation .products do not inherit their grain size from the austenite on cooling through the transformation range for does austenite inherit its grain size from the products which form austenite on heating. 5. In general, in both iron and carbon steel the larger the austenite ,grain size, the larger will be the grain size of the transformation products on cooling. This, of course, assumes all other conditions constant except the austenite grain size. An exception is found to this general rule in very pure iron such as electrolytic iron. 1n this instance small austenite grains may form very large ferrite grains on cooling through Ar3.

Jan 11, 1917

On Monday, October 25th, 1971 at approximately 12.15 a.m. the night shift proceeding underground discovered smoke at No.1 and No-3 cut throughs on the eastern side of the belt heading and in close proximity to the upcast shaft. Attempts to control the fire continued throughout the morning; as the position underground altered various procedures were adopted in an effort to contain and control the fire. Despite these efforts at approx. 12.40 p.m. a sudden change occurred which made it necessary to withdraw all workmen from underground, the vicinity of the Tunnel entrances and upcast shaft. Sealing of the mine at the entrances was the only remaining alternative. Several boreholes were put down to enable sampling of the atmosphere and air temperatures to be taken. This was essential as a guide to determining when conditions were suitable to allow exploration of the area using mine rescue self contained breathing apparatus.

Jan 1, 1972

By Robert G. Peets

After 216 years, the end of operations at the Corn- wall mine can be foreseen within the next two decades. The story starts in 1732 when three sons of Wiliam Penn-John, Thomas, and Richard-deeded 9669 acres in the Province of Pennsylvania to a Joseph Turner, who assigned it to William Allen. In 1734 Peter Grubb, who had prospected for iron in the area, paid Allen £ 135 for 300 acres of this tract in what is now Lebanon County. In 1737 he was granted a warrant for 142.5 additional acres, which included three hills of outcropping magnetite described as "rich and abundant, forty feet deep, commencing two feet under the earth's surface." This seemingly inexhaustible reserve was to become one of the most valuable deposits in the U. S. before the ore-rich Lake Superior region was developed.

Jan 7, 1957

By E. P. Pfleider, E. G. Malmlow, F. Landau

Friction drive hoists, long popular on the Continent, are now being installed in the U. S. and Canada by the Swedish firm of ASEA. Invented by C. F. Koepe and used first in 18.77, the principle is the same as that in elevator installations throughout the world. Inherent advantages of this system of hoisting were not fully developed until the recent introduction of the multiple-rope principle and design of concrete tower installations. Operators have a choice of independent operation of single skips and cages using counterweights, or use of skips or cages in balance. Also, hoisting control may be automatic, push button, or manual.

Jan 1, 1955

By Morton Handelman

The extraction of aluminum oxide_ from its ore is the initial step in the commerical production of metallic aluminum. This is most commonly done by the Bayer Process in which a caustic solution is used to dissolve hydrated alumina from the insoluable materials with which it is associate in bauxite. At the two-thousand ton per day Sherwin Alumina Plant of Reynolds Metals Company, the extraction residue from Caribbean ores is removed from the main pro¬cess stream as the underflow of a continuous clarifier-thickener operation. This sludge, prior to its ultimate r4jection from the process, is washed in a nine-step continuous countercurrent decantation system to recover physically-entrained mother liquor. Control of alumina production costs requires the optimum recovery of solubles from this residue which is generated at the same rate as the finished product. The equipment, operating techniques, controls and process variables which affect the efficiency of this soda recovery from Bayer Process red mud will be discussed.

Jan 1, 1959

By Nicholas Kyriazi

Significant portions of the present Federal regulations for certification of closed-circuit breathing apparatus (42 CFR 84) are not quantitative. The human-subject testing portion of the regulations specifies activities rather than metabolic work rates; however, heavy human subjects have higher oxygen consumption, CO2 production, and ventilation rates than lighter human subjects performing the same activities. This means that apparatus certified using different human subjects have effectively passed different tests. In addition, apparatus performance is monitored only during selected intervals of testing, and always during rest, rather than during the entire performance period. Further, testing is arbitrarily terminated at specific rated durations, leaving unknown the performance during the remaining duration. Duration itself is misleading to users who believe that the apparatus will last a certain time no matter how much they weigh or how hard they work. The actual duration obtained from an apparatus, however, is highly variable, being inversely proportional to the work rate at which it is used. In this report, recommendations are made to (1) replace the present regulations for certification of closed-circuit breathing apparatus with quantitative tests specifying metabolic work rate rather than activity,(2) continuously monitor relevant physiological stressors, (3) continue testing until exhaustion of the oxygen source, and (4) classify closed-circuit breathing apparatus by quantity of usable oxygen, i.e., capacity, rather than by duration. Physiology-based stressor levels are suggested as well.

Jan 1, 1999

By Suresh Menon

The explosive industry in India has passed through various stages - from a period of total dependency to a period of self sufficiency and now, is saddled with large spare capacity. The last decade and a half saw the emergence of small scale manufacturers, which until then was thriving on International technology.

Jan 1, 1998

COMPARED with the equipment used in most industrial processes, the open-hearth furnace has a relatively short life. The most important quality of an open-hearth refractory, therefore, is its rate of failure. This discussion of the physics and chemistry of the individual refractory materials will keep uppermost the question "Why, how, and where does it fail?" CAUSES OF FAILURE The principal causes of failure of open-hearth refractories are, in their order of importance : 1. Reaction with iron oxides. 2. Reaction with slag. 3. Flow and failure under load. 4. Melting. 5. Spalling. 6. Mechanical breakage. Reaction with Iron Oxides and Slag. Reaction with iron oxides and slag is by far the most important cause of failure. Liquid iron oxide comes near to being the "universal solvent" sought for by the alchemists. Among the materials commonly used in the open-hearth furnace, magnesia has the greatest resistance to this liquid and fireclay the least. The relations of iron and oxygen at high temperatures are illustrated in Fig 14-18. The liquid iron oxide that forms wherever iron is in contact with air or combustion gases cannot be given a definite formula. The percentage of oxygen at steelmaking temperatures may range from 21 to 28 (iron 79 to 72), depending upon the temperature and the oxygen pressure and also upon whether or not the liquid is in equilibrium with metallic iron. Oxide mixtures exist not only in the solid and liquid phases but also in the vapor phase, for metallic iron and its oxides have

Jan 1, 1964

By W. R. Ingalls

In choosing the unusual title given to this paper, in which the term "cyclic " might be substituted for "intermittent," my idea has been simply in respect of precision. We might say old method vs. new method, but that antithesis implies something of substitution that has not occurred and perhaps is not in sight. We may say horizontal retorts vs. vertical, but none of the older retorts are truly horizontal, and there are three different types of vertical retorts. Other complexities might be suggested. There is, however, a broad distinction between distilling that is accomplished in a periodic cycle, generally daily; and distilling that is accomplished continuously, irrespective of whether it be done in the vertical retorts of the New Jersey process or in the retorts of the Trollhät-tan process. In offering a summary of the present state of the art with the intermittent process, I shall not have any idea of exemplifying it through descriptions of any single plants, but shall rather divide among the main steps of the process and outline what is best developed among them. In pursuing this thought, it will manifestly be impossible to make a composition that will exemplify the acme of good practice, for the latter will always be a function of the cost of labor, fuel, etc., which are widely variable. As of 1929, in respect of labor we might generalize the average plant wage in the Pittsburgh-Wheeling district of the United States at $5 per day; in Oklahoma at $3; in Great Britain at $3; in Germany at $1.75-$2; in Belgium and France at $1. With high labor costs the metallurgist will naturally use means of the nature of mechanization that with low labor would be uneconomical when capital charges are taken into account. Nor may we from records of performance safely make positive deductions in respect of the efficiency of different types of furnaces unless they have been tested comparatively under identical conditions in the same works. We shall find gas-fired regenerative furnaces showing coal consumption from 1.2 tons down to 0.7 per ton of roasted ore, which may be ascribable to difference in the thermal value of the coal that is used or to differences in proportions of the parts of the furnace, which is something quite distinct from the principle of the furnace. Thus, we have

Jan 1, 1937

By T. D. Thomas

CONTENTS PAGE T. D. Thomas-Misfires in Anthracite Coal Mines 3 W. H. Forbes-Misfires in Bituminous Coal Mines 12 A. W. Worthington-Misfires in Non-metallic Mining (Limestone) 18 Misfires in Anthracite Coal Mines BY T. D. THOMAS,* LANSFORD, PA. (New York Meeting, February, 1929) IN THIS paper, major attention is given to misfires in mines where electric multiple shot-firing is the system used. Misfires are sometimes caused by one action or condition and at other times by a combination of two or more. They may be divided into two classes: (1) those caused by workers in the mines, (2) mechanical or manufacturer's causes. The second cause will be discussed first.

Jan 1, 1929

By Frank A. Peters

This description of process evaluation techniques used by the Bureau of Mines for studying proposed processing techniques has been prepared to provide those interested in using these studies with an understanding of the methodology. The methods used for preparing both capital and operating cost estimates are described. Each step in preparing the evaluations is described. Types of equipment costs are shown and factors used in the estimates are defined. Uses of the estimates are suggested and several examples are included. The accuracy of various types of estimates is discussed, along with an explanation of the reason for selecting the type used in the Bureau's process evaluation studies.

Jan 1, 1987

By J Silva Castro, W C. Wedding, B T. Lusk, J Calnan, E G. Morris

The emergence of electronic detonators for use in production mine blasting has enabled mining professionals to rethink the traditional blast design methodologies that pertain to timing. In many ways, electronic detonators can allow great advances in productivity, cost effectiveness, fragmentation, and safety. This paper reports on an independent study quantifying the accuracy of two electronic and two non-electric systems in which 674 detonators were tested. This research aimed to determine the accuracy of commercially available electronic detonators as compared to non-electric shocktube systems. Each system was tested over the viable ranges of delays available. Statistical models were then applied to quantify the scatter present in each system. Suggestions for the use of electronic detonators are made from the findings of this work. This work was made possible through funding from the United States Department of the Interior Office of Surface Mining Reclamation and Enforcement.

Jan 1, 2012

By David Miller

"The biggest concern about using soil mixing is the inability to observe physically the final product, which has led to a lack of confidence about what is in the ground. After all, you can’t see it and you can’t taste it and you can’t smell it, so are you sure it’s soil mixing? In addition, there are variations in techniques, tools and monitoring that have led to additional skepticism and the desire to “elevate” the product. This skepticism is why the industry is so scrutinized even when there are other technologies being used that do not experience nearly the same level of scrutiny. With other techniques, even with a slurry wall, you can see the backfill before it goes into the ground, and drilled piers have careful evaluation of the concrete mix even before placement into and verification of the open hole. Essentially, you get control over the final material. There is less control in soil mixing, and it’s very dependent on in-situ conditions. These factors make it difficult to rely solely upon installation parameters, quality control (QC) and verification tests. At some point, you may need to apply some engineering judgment.Start at the BeginningThe availability of sufficient and accurate geotechnical data (soil type, heterogeneity, shear strength parameters, variability, etc.) is important in establishing the material that will be drilled. This data will also help in determining the proper mix design to achieve the desired soil cement properties. Many factors must be considered when designing the mix and the mixing program, including any debris in the ground, different soil properties and varying moisture contents.Each of the various factors affect the final properties of the soil cement, so it really takes careful planning and evaluation of the mix design through iterations by looking at the soil types and adjusting the mixes. Clay is going to behave very differently than sand in both mixing efficiency and final properties. That is, the presence of clays results in more difficulty in achieving higher strengths and to physically mix the soil and the binder, whereas the presence of sands results in more difficulty controlling the final permeability."

Jan 1, 2017

By Garth M. Crosby, F. McIntosh Galbraith, Bronson Stringham

THE Hercules mine is located in the northeastern section of the Coeur d'Alene district, approximately 1 1/2 miles north of the town of Burke, Idaho. Surface indications of the ore deposit were first discovered in 1886, but regular mine production was not started until 1902 and was continuous until April 1925, when the known ore had been extracted. Incomplete records show that from 1912 until operations were suspended the mine produced 2 1/2 million tons of ore containing 9.4 pct lead and 7.7 oz of silver per ton, together with an estimated 2 pct zinc, 0.3 pct copper, and 20 pct iron. This operation was the first in. a series of mining enterprises culminating in October 1947 with the consolidation of Day Mines, Inc. In the same year it was decided to unwater the levels below the collar of the Hercules shaft in the hope of finding some indication of a recurrence of ore. The unwatering operation has been described in a. previous paper.' The initial exploration, following recapture of the workings, showed sufficient promise to warrant a detailed study of the mineralogy with modern techniques. The general geo1ogy of the Coeur d'Alene district, including a detailed description of the rock types encountered, has been comprehensively treated by Ransome and Calkins' in their classic paper, and only local background description, therefore, is felt to be appropriate here. The Hercules deposit transects a portion of the trough of a broad south-trending synclinorium which has been greatly complicated by faulting. More locally, it lies within a block of ground bounded on the east; by the O'Neil Gulch fault, a steep north-south overthrust of considerable magnitude, and on the west by a monzonite stock, the outcrop of which is 1/2 mile or more wide and 5 miles long. The country rock is composed of thin to medium-bedded argillites and argillaceous quartz-ites of the Prichard and Burke formations, the oldest members of the Pre-cambrian Belt Series of sediments in the area, believed to be of Algonkian age. The contact between them is a conformable gradation. The argillite is colored gray to tannish-gray and is fine-grained, compact, and generally massive in structure. Under the microscope the unaltered argillite is seen to be composed principally of anhedral quartz and a few feldspar grains which were at one time presumably partly rounded sand grains, but as a result of recrystallization and cementation by silica, the interstices are now almost obliterated and quartz grains show crenulate boundaries. The sizes of these crystals vary from 0.5 mm down to 0.1 mm in greatest dimension. In all specimens sericite comprises 10 to 20 pct of the rock and is present abundantly between most of the grains as flakes or shreds which vary considerably in size. Sometimes they form a fine felt-like mat or aggregate, and sometimes flakes are seen which appear to be good muscovite. In some specimens, separated rhombic-shaped carbonate grains are abundant, and in some instances these have been changed to sericite. Mining operations to date have explored the Hercules vein to a maximum vertical depth of 3600 ft below its outcrop, and along a maximum strike-length of 3600 ft on certain of the lower mine levels. The main orebody is irregular in outline, extending over a variable strike-length of 400 to 1500 ft; and it is intersected by a strong transverse fault that has been traced from the surface to the bottom level. This has been named the Hercules fault, and apart from the vein itself, it is the most prominent structural feature in the mine. There is good evidence that it existed prior to the introduction of ore solutions and may have influenced ore deposition, but it was also the locus of important post-ore displacement and shows a progressive right-handed horizontal component reaching 200 ft on the deeper levels. Its vertical component is not definitely known but may be considerably greater. The fault strikes 20° N to 50° E and dips westerly at angles of 70" to 45", flattening in dip where it crosses the original orebody from east to west between 1000 and 1600 ft below the surface. At about 3000 ft in depth the Hercules fault is joined by a vertical fault of similar strike, and the major post-ore dis-placement below their junction is taken up along this vertical branch of the structure, now called the Mercury fault. Recent work has been concentrated in this vicinity. Another structural feature of special geologic interest, though of little economic importance, is the occurrence of a porphyritic dike in this area. This lies a short distance above the Hercules fault, essentially parallel to it, and is 5 to 15 ft in thickness. It appears at first glance to cut the mineralization, suggesting push-apart relationship, but small stringers of the vein minerals have been observed to penetrate the dike for a matter of inches at several points. The dike is thought to be related to the monzonite intrusion. A vertical longitudinal projection of the mine is shown in Fig. 1, which illustrates most of the features discussed above. The Hercules vein was deposited along the course of a strong, persistent shear zone that now appears as a braided network of gouge seams running through more or less crushed and shattered country rock. It strikes 70° N to 80° W and dips southerly at an average of 75". Barren parts of the structure vary in width from less than 1 ft to more than 15 ft. The width of mineralized segments may be double that. Although the evidence is not conclusive, pre-mineral, normal movement along the zone may be 1000 or 1500 ft. The horizontal component is unknown. Post-ore movement appears to have been

Jan 1, 1954

By K. D. Jacob, W. L. Hill

THE bulk of natural phosphates is comprised of calcium phosphates, which are usually apatites;1 calcium aluminum phosphates such as pseudowavellite;2 and aluminum phosphates, which occur in extensive deposits, notably on Grand Connetable Island.3 Nearly all deposits of current commercial value are apatite, often with admixed material of other classes. In the trade coarsely crystalline material, mainly fluorapatite of igneous origin, moves as apatite, whereas the fine-grained material of sedimentary origin is marketed as phosphate rock. The latter classification includes the phosphate shales in western North America. Although several genetic varieties of phosphate rock are recognized,4 most known reserves are of marine origin. Calcium phosphate is concentrated from Florida land pebble and Florida hard rock deposits and from Tennessee brown rock deposits. Phosphate Rock as a Fluorine Carrier: More than two decades ago Reynolds and Jacob surveyed the fluorine content of commercial phosphate rock from various parts of the world.3,5 Later these samples were reanalyzed by a new and more reliable method of determining flourine.6,7 The revised results with interim additions of new samples, published a few years ago in a brief summary,' are shown in Figs. 1-3. Through the points representing the respective varieties of rock medial straight lines are drawn. In some instances the lines are well defined by the points; in others, where scattering renders the position of the best line uncertain, the greater weight is given points near the end of the covered range. Wherever distinct varieties are not recognized among regional deposits the custom is to identify the sample with the geographical location of the deposit.

Jan 10, 1954

INTRODUCTION This chapter summarizes characteristics of porphyry copper deposits within the Cordilleran orogen east of the Coast Range plutonic complex of the Yukon and British Columbia and south to the Columbia River plateau and the Idaho batholith. Porphyry copper deposits in this area reflect a geologic environment different from that found for mineralized intrusions elsewhere in the Cordilleran orogen, both to the south in the United States and Mexico and to the west in Alaska. A section on the Tertiary Cascade porphyry copper province is included herein. The Cascade province is considered separately because its deposits display a unique mineralogy and age of mineralization. That portion of the Cordilleran orogen described in this chapter is shown in Fig. 32, with the Cascades shown in Fig. 39 (see p. 114). The regional geology of this area has been described by Monger, et al. (1972), Hollister (1974), Wolfhard and Ney (1974), Field, et al. (1974), and Stacey (1974). Each offer slightly differing views on the post-Mississippian evolution of the geology in this porphyry copper province. The descriptive geology in this chapter has benefited from many discussions with J. E. Armstrong and J. W. H. Monger. Two separate petrologic models have been used popularly to categorize porphyry copper deposits in this area, a diorite and a quartz monzonite model. The diorite model, which involves a mineralized quartz-free pluton, has also been called the alkalic type (CZM Special Volume 15, 1976, and Soregaroli, 1975) and the syenite suite (Sutherland Brown, et al., 1971). The quartz monzonite model has also been named the calcalkalic type (CZM Special Volume 15, 1976) and includes mineralized quartz-bearing plutons compositionally ranging from quartz diorite to granite, with the igneous phase closest to ore in time and space commonly being quartz monzonite.

Jan 1, 1978

By Mohan Ram

Structural impurities in the crystal lattice of rock phosphate of Indian origin is a great challenge for beneficiation and practically possesses remote scope of upgradation to the desired level for industrial utilization. There is a wide occurrence of rock phosphate in different parts of India with different characteristics viz. siliceous, carbonaceous and ferruginous. The structural cationic replacement in the composition of apatite is due to the exchangeable cations present in the external impurities associated with the rock phosphate ore. The Indian rock phosphate ore from Purulia district, West Bengal is characterized as unusual structural composition due to the exchangeable cation in its composition. The elemental distribution of ore exhibits the disproportionate ratio of P2O5 and CaO which is 5.88: 3.31 in reference to the theoretical composition of P2O5: CaO in apatite which is 4.23: 5.55. The substitution by trivalent Al+++ cations within the structure of apatite in place of divalent Ca++ is found upto 18% and substitution of exchangeable cation in place of PO4 and F ion by CO3/SO4 is found upto 0.1% and OH ions upto 0.2% respectively in a lattice of apatite thereby diluting the grade of rock phosphate and restricting its industrial use even after beneficiation. This paper reveals the unusual compositional features of rock phosphate under study and is a guideline to the mineral industries using such rock phosphate. Keywords: external impurities, exchangeable cation substitution, structural compositional features, electron probe micro analyzer

Sep 1, 2012

By M. I. Brittan

Discussion Dr R. E. Robinson (Fellow): The author must be congratulated on a very meticulous and self-contained piece of work. It is indeed a pleasure to read a paper that is so clearly and systematically laid out, and where the conclusions and the testwork conducted have been so clearly described. The paper is complete in itself, which makes it very difficult for someone who is not intimately involved in the whole Torco project to make any comments on its content. However, two points that, strictly speaking, fall outside the scope of the paper are of considerable interest. The first relates to the particle size of the material treated. In the paper, the testwork is confined to one standard particle size (minus 60 plus 100 mesh). The essential feature of the paper is to indicate that the rate-controlling reaction in the whole segregation process is the rate of reaction of the ore particles with the reducing agent and the hydrogen chloride. One wonders, therefore, to what extent this relatively slow rate of reaction is affected by the particle size of the ore itself. One imagines that the reaction must take place by contact of the hydrogen chloride with the surface of the mineral particles, and it is reasonable to suspect that the rate of diffusion of the copper ions to the surface is a relatively slow process and is thus the limiting factor in this particular rate of reaction. It is possible, for example, that the improvement obtained, when the ore is subjected to reducing conditions before the chlorination, is due to a breakdown in the crystal structure of the original particle. This breakdown is brought about by the reduction and by the consequent increase in surface area available for reaction with hydrogen chloride. Can the author indicate whether any work has been done along these lines, and whether it has been established that the reaction depends on the surface area available? The second point relates to the application of this kinetic study to the actual operation of a Torco reactor. It was once planned to feed the sodium chloride, together with the reducing agent, into the top of the segregation chamber. In the paper, the author mentions that it has now been established that the segregation chamber behaves, to all intents and purposes, as a fluidized bed, and that there is, therefore, a rapid evolution of gas in the lower regions of the chamber, which, it is imagined, displaces the gas phase rapidly. Since the reaction between sodium chloride, water vapour, and the aluminium silicates in the ore is extremely rapid, one wonders how much of the hydrogen chloride produced is removed from the reaction zone before it has had time to react with the copper minerals. The extremely low consumption of sodium chloride (which is a vital feature of the Torco process) must depend on an extremely rapid circulation of the hydrogen chloride gas to all the ore particles in the segregation chamber. One wonders, therefore, if a system for the introduction of the sodium chloride into the bottom regions of the chamber might not result in even greater efficiency in the utilization of sodium chloride. ProC D. D. Howat (Fellow): All of us who have been concerned with the study of chemical reactions at high temperatures are keenly interested in kinetics and are well aware that this is not an easy study experimentally. Dr. Brittan is to be congratulated on the development of neat experimental methods and for his full discussion of the results obtained. Although the segregation process for the extraction of copper from oxide and silicate ores has been known for almost fifty years, the fundamental chemical and physical changes involved have been little understood and the fundamental data are very scanty. The work now in progress at A.A.R.L., together with that sponsored by the Anglo American Corporation in other research institutions throughout the world, is bound to produce new fundamental data and a much more complete understanding of this rather fascinating process. It is already apparent that some of the old and well-worn chemical reactions that were postulated to occur, just cannot take place in the way which was formerly accepted. Dr Brittan's work, carefully conducted and thoroughly analyzed as it has been, still leaves us with one great outstanding problem. The thermodynamic data and the possible reactions set out in Table I (page 281) of his paper leave us asking, in complete despair, how can copper be converted into a volatile chloride in the presence of HCl, CO and carbon at temperatures about 800°C? The thermodynamics all combine to show that copper should be reduced to metal as the first step in the process. This brings us right up against the second problem. If copper were reduced to the metal how would HCI convert it to the volatile chloride? On top of these problems is the unknown reason for the very high speed of reaction between CO, HCI and the ground copper ore. Still further into the region of the unknown is the reaction by which gaseous hydrochloric acid is produced in the actual process. Perhaps Dr Brittan is feeling grateful that he doesn't have to try to explain this reaction-at this stage of the research programme at least. The results very clearly show that both CO and HCl gas are essential for rapid production of the volatile copper chloride. Dr Brittan states that 18 minutes were required to attain 83 per cent extraction with HCl gas alone and this was reduced to 4 minutes when CO was

Jan 2, 1970