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By H. R. Gardner

THE wide variation in tension and compression properties' which exist between the a,ß and ? phases of plutonium indicate that a study of the impact properties would give a greater insight into the variables affecting the mechanical properties and formability of plutonium. In tension the a phase is quite brittle with elongations less than 1 pct. The ß phase at normal test speeds is unusually ductile; elongations up to 621 pct have been obtained. However, the ß phase is quite sensitive to strain rate; elongations decrease rapidly as strain rate is increased. The y and 6 phases are less ductile than the ß phase, with the 5 phase the more ductile of the two. Neither are very sensitive to strain rate. Considering the brittle nature of a Pu in tension and compression and the strain rate sensitivity of ß -phase Pu, both unnotched Charpy specimens and standard V-notched Charpy specimens were fabricated for testing. In addition, since iron in amounts greater than 500 ppm forms a Pu-PlgFe eutectic network which inhibits microcracking in as-cast plutonium,' it was decided to prepare impact specimens with low and high iron contents. Chemical analysis and density data for the impact specimens are presented in Table I. A total of 55 impact specimens, 32 unnotched and 23 with the standard Charpy V-notch were tested at temperatures from —45' to 330°C. The test results for both unnotched and notched specimens are illustrated in Fig. 1.

Jan 1, 1962

By Paul S. Parobeck, Robert A. Haney, Andrew J. Gero, Thomas F. Tomb

As a result of recommendations made in the 1988 Report of the Mine Safety and Health Administration Advisory Committee on Standards and Regulations for Diesel-Powered Equipment in Underground Coal Mines, the Mine Safety and Health Administration (MSHA) initiated a program to develop and test a diesel particulate sampling device. The sampling device employed the principles of inertial impaction to separate respirable dust into two fractions; one fraction containing particulate material having an aerodynamic equivalent diameter equal to, or greater than, 1.0 micrometer, and the other fraction containing particulate material having an aerodynamic equivalent diameter less than 1.0 micrometer. The design criteria were based on data which showed that virtually all of the diesel particulate material and very little of the mine generated dust was less than 1.0 micrometer in size. The instrument was tested in underground dieselized coal mines and in the laboratory. Comparative measurements were obtained with the MSHA designed impactor sampler, a University of Minnesota (UM) designed impactor sampler and MSHA's standard respirable coal mine dust sampler. Simultaneous measurements were also obtained with five of the MSHA and five of the UM designed impactor samplers to investigate between sampler precision. The paper discusses the design differences between the MSHA and UM impactor samplers, the precision of replicate measurement obtained with both instruments and compares measurements obtained with the respective samplers of the sub- and supermicron fractions of the respirable material.

Jan 1, 1991

By F. W. Robinson

THE temperature coefficient of electrical resistance of pure metals is high and therefore the resistance increases rapidly with rising temperature. In 1871, Siemens suggested the use of this property as an accurate means of temperature determination. Owing to practical difficulties, particularly the contamination of the metal and consequent permanent change in its resistance and its resistance-temperature relation, this method fell into disrepute as a practical standard until revived later by Callendar and Griffiths, and subsequently by Holborn and Wien, all of whom showed that the earlier difficulties were not inherent in the method but incident to the mode of protection of the resistance coils. Following .the work of these investigators, the problem of temperature measurement by this means has been the subject of careful study and has now assumed an importance second only in practical adoption to the thermo-electric method. It is generally recognized, both here and in Europe, that the standard temperature scale should be the thermodynamic as it permits the evaluation of high temperatures on the basis of the radiation laws of Stefan and Bolzmann, of Rayleigh, and of Wien and Planck, on a scale consistent with that obtained by means of the gas thermometer at low temperatures. Lord Kelvin showed that only. a very small correction, amounting to about +0.7° at 1000° C.-almost within the limits of experimental errors-is necessary to adjust the constant-volume nitrogen thermometer to the ideal thermodynamic scale.

Jan 9, 1919

To the members of The South African Institute of Mining and Metallurgy I have audited the annual financial statements of The South African Institute of Mining and Metallurgy set out on pages 641 to 644 for the year ended 30 June 2006. These annual financial statements are the responsibility of the Institute?s Council members. My responsibility is to express an opinion on these annual financial statements based on my audit. Scope I conducted my audit in accordance with International Standards on Auditing. Those standards require that I plan and perform the audit to obtain reasonable assurance about whether the annual financial statements are free of material misstatement. An audit includes: ?examining, on a test basis, evidence supporting the amounts and disclosures in the annual financial statements, ?assessing the accounting policies used and significant estimates made by management, and ?evaluating the overall financial statement presentation. I believe that my audit provides a reasonable basis for my opinion. Audit opinion In my opinion, the annual financial statements present fairly, in all material respects, the financial position of the Institute at 30 June 2006 and the results of its operations and cash flows for the year then ended in accordance with generally accepted accounting practice appropriate to the Institute. Supplementary information The supplementary schedule does not form part of the annual financial statements and is presented as additional information. I have not audited this schedule and accordingly do not express an opinion on it.

Jan 1, 2006

Discussion of the paper of E. J. Hall and C. W. Drury, presented at the Butte meeting, August, 1913, and printed in Bulletin No. 78, June, 1913, pp. 1059 to 1066. A. M. SMOOT,* New York, N. Y. (communication to the Secretary +) :-The nails method is given in many text books on assaying as a standard method for gold or for gold and silver in ores. It appears to be taught in the laboratories of technical and scientific schools as a standard method, and to some extent it is used by assayers in actual practice, although the class of ores to which it may be applied is limited to those free from reducible metallic compounds, the metals of which would enter the lead button. The experiments of Messrs. Hall and Drury show the inaccuracy of this rnethod for silver, in heavy sulphides, but not so clearly the short-comings with regard to gold. It seems certain that low silver is caused by the solubility of silver in iron-alkali sulphide as compared with its solubility in molten lead; that is, silver is distributed between iron-alkali sulphide and metal in the charge in proportion to solubility at the particular temperature of the fusion. Lead is a fixed quantity in these assay charges, but iron-alkali sulphide increases with increased sulphur, so the more sulphur in the charge, the lower the silver results.

Jan 11, 1913

By R. V. Ramani, V. T. Burgos, J. A. McClay

The systems contributing to the cost analysis, and the flow of data input and output related to these systems, are brought into focus in Figure 6. The various sub-systems of the Master Design Simulator provide the raw operating data basic to the cost model which, operating independently, accumulates the raw input, assigns costs to this data by standard or modified costing procedures, manipulates the costs into standard formats and outputs the data for managerial evaluation. Cost Modeling Direct costing methods are not complex in the sense that elemen¬tary arithmetic is employed using the same basic relations for each cost category. The basic cost model relations described above are illustrated in Figure 7. The greatest difficulty encountered in the transition from a dy¬namic systems simulation model to a costing model is the time frame of the input and output data. Dynamic simulation models utilize a small time increment, on the order of seconds or fractions of seconds, to analyze system operations and interactions. On the other hand, for cost computation and reporting, the smallest relevant time unit is the shift, with reporting periods expanding to days, weeks, months, and years. Complications arise in the accumulation of data for the various report periods due to the use of the Gregorian calendar by modern so¬ciety. The potential for simulation for the life of the mine, possibly in excess of 50 years, increases the complexity of the costing process

Jan 1, 1975

One of the most frequently asked questions at the feasibility study stage of a project is, æhow big can one go?Æ This important question effects costing and efficiency of metallurgical equipment, the real estate required and the operating plant strategy.The presentation covers both engineering and process considerations and explains the limits that exist today and explores future trends.The paper is specific about large slurry application usually found in gold applications; however, it could apply to silver and uranium or other dense media leach duties.Agitator drive sizes are examined as well as process risks associated with large tank agitation. New technology has increased the maximum tank size and designers are now experienced in tank volumes in excess of 5000 m3. Tanks of 6000 m3 have been agitated and 10 000 m3 tanks are within present engineering know-how and scale-up.Fabrication techniques for shaft and impeller designs exceed the existing standards and new specifications will be needed to cover large mixer design. The paper covers certain aspects of mechanical design and proposes alternative standards that are more appropriate. Process considerations such as fluid natural frequency as well as typical superficial velocities for efficient operations are examined. Scale-up techniques based on existing and new technology are also examined to ensure a successful installation.The presentation therefore covers the majority of large mixer design questions that are frequently asked.

Jan 1, 2006

By J. L. Gaigher

A simple mineralogical regression model is described, which was devised in order to help the flotation plant cope with rapidly changing headgrades and feed mineralogy. Minerals in the feed were determined by means of quantitative X-ray diffraction and compared with plant performance parameters and flotation reagent consumption rates, using regression analysis. A mineralogical regression model for predicting recovery served to motivate flotation plant personnel and also set standards whereby flotation performance could be measured. The reagent consumption model indicated in which direction any or all of five flotation consumption rates should be altered. All practically measurable parameters, even those showing poor or no correlation, were included in the model, provision being made for the continuous update of a batch of data of 30 lines. Compared with the method of multiple linear regressions, the r2 is poorer at 0,41 (versus 0,50), but the mean standard error of estimate for both methods is identical at ± 6,6 units. The regression model is simpler than the method of multiple linear regressions, and although the accuracy of prediction is not high (41%), the mineralogical regression model proved to be successful in practice. Flotation efficiency, as measured by the mean ratio of head to tails grade, for the 9 month period after the mineralogical regression model was implemented on an hourly basis, improved from 2,8 to 3,2, which is equivalent to an improvement in average flotation recovery of 5% relative.

Jan 1, 1987

By Alan Campoli

Approximately 100 million fully grouted rebar roof bolts are installed each year in the United States, 80% are B-Series (5/8 inch fully grouted rebar in a one-inch borehole). The lesser used more expensive A-Series (3/4-inch fully grouted rebar in a one-inch borehole) maximizes resin performance. Standard B-Series provides only 60% of the pull out resistance of A-Series. The Eclipse System was developed by Minova USA Inc. and marketed with Excel Mining Systems to improve the performance of B-Series. A 1/8-inch offset bolt head combined with a specially formulated Eclipse Lokset resin cartridge results in increased mixing efficiency and more uniform resin anchorage. Over two million units have been sold to date; the majority of these Eclipse bolts have been sold to coal mines in the Illinois Basin. Pull testing has demonstrated that a Grade 60 Eclipse bolt provides equal or superior pullout resistance as compared to a Grade 40 No. 6 rebar. Short encapsulation pull tests have yet to consistently demonstrate the superior pullout strength of Eclipse to standard B-Series, however 12 inch Eclipse grouted columns have successfully held in excess of the bolt yield load. The successful substitution of Eclipse bolts for A-Series has significantly reduced the primary roof support costs by approximately $0.18 per foot of bolt. Recent over coring analysis of various anti-gloving methodologies performed in New Zealand has confirmed the superior mixing of the Eclipse system (5).

Jan 1, 2004

By S. K. Murphy

A partial shaft pillar extraction project was embarked on at a mine in the West Wits mining region in South Africa. Prior to the extraction, a rock engineering assessment was conducted with the main objective being to establish which areas of the shaft pillar could be extracted without unduly affecting the stability of major infrastructure within the shaft pillar. Standard numerical modelling methods were used to conduct this assessment and several areas were identified for extraction. However, during the extraction process, it was found that the infrastructure was affected more than was originally predicted by linear elastic modelling. In addition to this, several large events resulted in far more large aftershocks than would normally be predicted. In order to gain an understanding of this phenomenon, seismic data was integrated with numerical modelling as follows: > The seismic data in the form of displacements was integrated in 3D space within the model to assess the effect of static stress changes associated with seismic activity > Data from large events was integrated onto specific structures, where displacements associated with large events were integrated onto the plane of structure in question. This paper describes the methodology and results obtained from each of the above processes and include insights into the mining process that were previously unfathomable using standard linear elastic numerical modelling. The way forward for this rock engineering assessment of partial shaft pillar extraction is also discussed.

Jan 1, 2012

Standard petrographic profiles for many Permian seams from New South Wales and Queensland have been drawn up. The standard profiles of seams of the same age and from one or more coalfields were used to compile an idealized or type seam sequence for those coals in which more than half the seam profiles were similar to one another. A marked similarity existed between at least 7 of the 10 type seam sequences formed. The configurations of the vitrite-plus-clarite contents of the plies of 95 per cent of the seams could be formed from different arrangements of 3 basic sequences: (i) a two-ply sequence with a lower vitrite-rich and an upper vitrite-poor ply; (ii) a two-ply sequence with a lower vitrite-poor and an upper vitrite-rich ply; and (iii) a four-ply sequence of 3 plies with vitrite-plus- clarite decreasing upwards, terminating in a vitrite- rich ply. The third basic sequence was very similar to a succession described for some British Carboniferous coals. Eighty per cent of the Australian Permian coal seams studied began with the first (i) basic type of sequence and are interpreted as originating in peat- forming areas which became progressively drier. More than half the seams were formed in an area with a stable depositional surface level which was subject to oscillations of small amplitude.. Many (30%) were formed in an area with a steadily rising depositional surface level.

Jan 1, 1969

By Burton Scott

A device for detecting carbon monoxide at ppm levels is described which is based on the measurement of infrared absorption at 4.65 microns. The unique feature of the nondispersive infrared instrument is contained in the optical design which includes a confocal cavity, increasing the effective optical path length by a large factor, thus greatly reducing the required package size. The cavity has the effect of reducing the total amount of radiation passing to the detector, but at the same time increasing the percent attenuation observed when adding CO, by a factor called gain. The radiation is chopped by a filter wheel which contains narrow band filters at 4.0 and 4.65 microns. The resultant signal produced by the pyroelectric detector is demodulated by the electronics to product an automatic gain control signal, proportional to the total radiation, and a difference signal, proportional to the CO present. Tests of a breadboard model are described which show cavity gains of 10 with noise levels equivalent to 2 to 10 ppm CO.

Jan 1, 1975

A drive, 3,500 ft long, was planned to explore mineralization exposed by diamond drilling. A rate of advance of 150 ft per fortnight, approximately double previous Broken Hill rates, was required. Subsequent to the completion of this drive its extension for a further 6,200 ft was planned to provide a main haulage way along which ore could be transported from a newly acquired lease back to the main hoisting shaft, a total distance of 13,500 ft.This paper describes the working conditions, relationships, manning and costs, and gives details of the cycle and equipment used to achieve rates up to 170 ft per fortnight.OBJECTIVEDiamond drilling from the 1,000 ft level at the Delprat Shaft section had indicated that the zone of mineralization, about 150 ft wide, which had been explored on the 1,000 ft level for a length of 2,600 ft, continued downwards with a westerly dip. The drilling results showed that within this wide zone there were arrower lenses of mineralization of possibly payable grade (see composite plan, Fig. 1).Diamond drill holes of 2,000 ft and over, particularly through heavily sheared areas of sericite schist, had been found to be costly, slow, and inadequate in testing the potential and trends of these smaller concentrations of mineralization.It was therefore decided that an exploratory drive should be advanced from the northern limit of the 1,480 ft level into the western mineralization, and that from this drive diamond drill holes should be drilled at dose intervals to outline the ore.The 1,480 ft level was chosen for the following reasons.1. It had the greatest northerly reach of the accessible levels at No. 7 Shaft section below the 1,000 ft horizon and therefore would enable the test to be made most quickly.

Jan 1, 1965

By T. H. Jeffers

The Bureau of Mines conducted a laboratory investigation of copper electrowinning from electrolytes produced by solvent extraction. The purpose of the research was to gain a better understanding of the fundamental relationship between copper electrowinning conditions, anode corrosion, and cathode purity. Processing variables were evaluated in a cyclic semicontinuous system in which the depleted electrolyte was used to strip copper from loaded organic extractant. For the production of high quality copper cathodes, the electrolyte was dosed with cobalt and stripped of organic solvent extraction reagent. An optimum cobalt addition of 60 ppm, and removal of the entrained extractant with activated carbon provided the best conditions for electrowinning copper containing the least amount of lead and controlling anode corrosion. As little as 13 ppm of the entrained organic extractant, LIX 64-N, significantly accelerated anode corro-sion. However, the addition of 1,000 ppm of organic diluent kerosine did not affect anode corrosion. After cobalt dosing and removing the entrained extractant, cathodes containing less than 2 ppm lead were consistently electrowon at widely varied current densities, electrolyte copper and acid concentrations, and temperatures. Dosing of the electrolyte with selected impurities including magnesium, aluminum, manganese, and iron did not affect the cathode purity. However, 0.05 giL of chloride ions markedly increased anode corrosion and the sub-sequent transfer of lead to the cathode copper.

Jan 1, 1985

By Aurel Goodwin

MSHA is the regulatory Agency which administers the Federal Mine Safety and Health Act. Although MSHA's principles largely derive from this Act, they do not exculsively derive from it. One might expect that our principles are contained in the regulations we have published to implement the Act. To a degree, this is true; but the regulations we have now were developed and published before the current Act became effective. For this reason, as well as others, there are some principles that we support which are not evident in our regulations, or in the Act. One of these is the ALARA principle. This principle requires that all unnecessary exposure be avoided; that is, all exposures should be kept "As Low As Reasonably Achievable." Almost all professionals in radiation control can agree with this broad principle; but it would be difficult to get agreement on a regulation which would translate the principle into industry practice. Rather, our regulations specify a limit to exposure for any individual. For radon daughters the current limit is four working level months per year. By having an explicit limit such as this, the ALARA principle often becomes lost and the limit becomes the goal. The principles contained in the Act are equally broad in scope. The principles apply not only for radiation protection, but also for toxic substances and harmful physical agents. The Act states that standards for such substances or agents shall most adequately assure on the basis of the best available evidence that no miner will suffer material impairment of health or functional capacity, even if such miner has regular exposure to the hazards dealt with for the period of his working life. The Act also provides that, although protection of health must be our foremost concern, other considerations shall be the latest scientific data in the field and the feasibility of the standards. Our regulations reflected similar principles when they were developed and promulgated. From experience with other health and safety laws, Congress realized that setting and meeting an exposure limit may not be sufficient to prevent disease. The Act, therefore, contains a detailed description of additional provisions to be included, where appropriate, in mandatory safety and health standards. One of these additional provisions requires that miners be informed about the nature of the hazards associated with their job and about the means for their own protection. Another provision requires the use of labels and other forms of warning to inform miners about the hazards, about proper precautions for safe use or exposure, about relevant symptoms of overexposure, and about emergency treatment. We have partially implemented these provisions through regulation. The Act requires also that standards shall prescribe protective equipment and control of technological procedures and that they shall provide for the monitoring of miners' exposures. We have implemented this provision to a degree also through regulation. Finally, the Act specifies that, where appropriate, a mandatory standard must prescribe the type and frequency of medical examinations in order to most effectively determine whether the health of miners is adversely affected by exposure. Additionally, when a determination is made that a miner may suffer material impairment of health or functional capacity by reason of exposure, that miner must be removed from such exposure and reassigned. In order to encourage miners to take the medical examination, the Act also provides that the miner shall not suffer loss of pay as a result of being reassigned. This additional provision on medical examinations has not been implemented by regulation for radiation hazards, nor for most other toxic substances or hazardous physical agents. We believe that this conference will provide us with valuable information on medical examinations for radiation exposure, as well as on other considerations to be used in future regulations. We still question whether our basic exposure standard of four working level months per year is adequate to protect a miner's health. This conference is indeed timely because both MSHA and NIOSH have been reviewing this issue for some time. I would like to reemphasize MSHA's commitment to education and training. MSHA strongly believes that miners should be informed fully about the real and potential hazards associated with their work. They should know the nature of the hazard and the means for protecting themselves from the hazard. The Act also emphasizes the need for training miners by requiring that new miners receive 40 hours of training before working underground and 24 hours of training before working on the surface. The Act also requires eight hours of refresher training annually. This training must include information on the relevant health hazards and the potential consequences of overexposure. Before I close, I would like to mention briefly another source of principles that MSHA often consults to obtain guidance for both the development and enforcement of regulations. These are the various court decisions that are handed down from time to time. Two in particular that I want to stress are the recent Supreme Court Decisions on benzene and cotton dust. Although neither of the decisions has a direct impact on mining or radiation, they do clarify some general principles and directions for regulatory agencies, such as the role of cost-benefit analysis and the need for risk assessment. In closing, I wish us all an open exchange of information for a productive and useful conference.

Jan 1, 1981

Due to highly competitive international coal markets, Stockton mine is bound to produce very high quality coals consistently if it is to maintain and expand its market share. The nature of the deposit is such that coals from various source pits are blended in differing proportions to meet the various coal quality specifications of the customer. The complexity arises due to quality variation within the seam both horizontally and vertically and the fact that up to six different products with completely different quality specifications may have to be produced. With increasing demand for this rare product and with CoalCorp eager to meet this demand comes the requirement for computer methods to handle this highly complex problem of both modelling and scheduling.

Jan 1, 1989

By David Siu

Many problems were encountered with the mechanical wedge-type splices used for the 15-inch square precast prestressed concrete piles installed at an industrial plant site. Many of the wedges were breaking during installation of the splices and after pile driving resumed following splicing. Dynamic monitoring of the piles and uplift loading tests confirmed that the problems were associated with cracked or damaged wedges. More ductile, annealed wedges were subsequently used, and, while not completely satisfactory, they were found to perform better than the original more brittle wedges. Introduction Long precast prestressed concrete piles were installed to support heavy, settlement-sensitive structures at an industrial plant in Port Alberni, British Columbia, Canada. The piles were driven through soft compressible silts and clays to toe bearing in a dense sand and gravel deposit or on bedrock. Each pile consisted of two or three segments joined with mechanical wedge-type splices trademarked DYN-A-SPLICE. Shortly after construction began, problems arose during pile installation, and these problems were, later, identified to be related to the wedges, or clamps, used to connect the splice plates.

Jan 1, 1991

By B. D. Paphane

Since the commissioning of the Tati Hydrometallurgical Demonstration Plant (HDP), which produces Ni and Cu cathodes as well as Co salt from sulphide concentrate employing Activox® technology for leaching, no kinetic data has been generated from the plant under normal operating conditions, from start-up to steady state. Therefore, this study seeks to determine the rate of reactions in the leaching circuit with a view to obtaining kinetic data for the system under the current operating conditions. Within the five-compartment autoclave operating at temperatures of 105°C to 110°C and 1100 kPa pressure, the extraction of Ni, Cu, and Co ions into solution is achieved through the oxidation of the sulphide concentrate using approximately 50 kg H2SO4 per ton, 10 kg NaCl per ton, and 1.21 kg O2 per kilogram of S2-. The autoclave reactions showed complex kinetics, with the reactions proceeding via some intermediate(s) before the formation of products. The results suggest that the reactions were dependent on the size of the metal ion of interest and the concentrations of the metal ions in the Tati concentrate as well as in chloride leaching. The order of the leaching rate constants was Fe2+ > Cu2+ > Ni2+ > Co2+ in all the compartments studied. The rates of leaching were also found to decrease as the concentrate moves from compartment 1 to compartment 4. For compartments 1 to 3, leaching was observed to be preceded by dilution of the repulp liquor (copper raffinate). The leaching rate for compartment 5 was slightly lower than that of compartment 1 but higher than the rates of compartments 2, 3, and 4 for all the metals studied. Keywords kinetics, hydrometallurgical plant, Activox

Jun 1, 2013

By A. Vien, B. C. Flintoff, R. P. Edwards

"This paper provides a critical review of the development and application of advanced controls in mineral processing at the industrial level. These algorithms can be sub-divided into four broad categories: Kalman filter, self-tuning, extended horizon, and knowledge based or expert systems. Applications from each category are described outlining the benefits derived, the relative advantages of that particular controller, and potential limitations of the controller which may inhibit their more widespread application. The paper will conclude with a description of control algorithms which are currently in the rudimentary design stage, but may hold tremendous potential for mineral processing control systems in the next decade. INTRODUCTIONThe Canadian mining industry has a large installed milling base and many plants are run with manual or low level automatic process control. The competitive pressures imposed by the market will not allow these operations to ignore the significant increase in efficiency that can be achieved through the effective application of advanced process control. These improvements will come in the form of better quality, higher productivity and reduced costs.There have been quite a number of reports of successful applications of advanced control in mineral processing applications. As this trend continues, and more of the work is documented, the technology will be more effectively transferred to process and control engineers. It is the authors' opinion that the most difficult problems facing process control are not really technical in nature. Rather, they lie more in process control education. This extends from management, through engineering to the operators.The purpose of this paper is to provide a tutorial type overview of some of the more frequently encountered advanced control algorithms. The assumption is that the reader has a superficial knowledge of this technology. In the discussions which follow, several techniques are explored within the framework of simple process examples. Industrial examples are cited for each and an evaluation of the expected longer term potential is given."

Jan 1, 1992

By B. M. Larsen, T. E. Brower

IN two previous papersl,2 dealing with the carbon-oxygen reaction, and the simultaneous content of each, in liquid steel in the furnace, we have made use of the quantity 0[O], defined as the excess oxygen beyond what would be in equilibrium with carbon then present if the pressure of carbon monoxide were I atm,; that is, if [C] X [O] is taken as 0.00222, 0[O] = [O] - 0,00222/[C] where [O] is the percentage of oxygen by weight as measured by the method described earlier.' This quantity 0[O] proves to be substantially independent of carbon content over the range 0.08 to 1.2 per cent, as well as of the basicity of the slag, and is nearly always within the limits 0.015 to 0.025 per cent when the bath is boiling normally and is in what we have termed a steady state. Apparently it must be at least about 0.014 per cent before boiling occurs; but this threshold value seems to depend upon the ease of bubble formation at the metal-hearth interface, hence upon the condition of the surface of the hearth. In any case, 0[O], on the average, changes considerably following any disturbance of the bath, as by addition of ore. The present paper deals with the average changes in ?[O] level following addition of Spiegel or crop ends, or after procedures to promote stirring, or by tapping of the steel from the furnace. The observations were made on a large number of commercial heats, and consequently are subject to the influence of various disturbing factors; the results have the compensating merit of showing the consistency obtainable under the limitations imposed by practical operation. VARIABILITY OF EXCESS OXYGEN IN NORMAL OPEN-HEARTH OPERATION The preceding general statements refer to the average value of ?[O] particularly in heats that are boiling actively; on individual heats, which are less active, there may be a variable time lag between the disturbance and its effect, which complicates the interpretation of what is going on. For instance, if in the normal refining process boiling is sluggish, the slag usually is gaining oxygen from the gas phase above it, or from added ore, faster than it is losing oxygen to the metal; yet when the bath starts to boil vigorously, the opposite condition may prevail for some time. The consequence of this lag is a variation of ?[O], the interpretation of which in a single heat may not be immediately obvious, This condition is illustrated in Figs. I and 2. In Fig, I, ?[O] at the beginning of the test period was near the lower limit of its usual range (dashed lines) in the steady state as a consequence of the lime boil, which had ended about as indicated. The bath is now cold, carbon drop is slow, and the first small ore feed causes only a small rise in 0[O]; as the temperature rises and as lime dissolves in the slag, and the bath becomes more active, the

Jan 1, 1946