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By Jun Lu, Min Chen, Wei-jian Su, Jian-zhong Luo, Yu-tao Lei

"In the case of treating phosphatizing wastewater containing zinc, zinc had an impact on the solution chemistry, especially on pH. Fe2(SO4)3 and CaCl2 were used as precipitators to remove phosphate from zincic phosphatizing wastewater and the optimum pH was investigated. Zinc ion had a significant effect on the optimum pH. The optimum pH ranges were 12.10(inlet concentration of 15.67mg/L, outlet concentration of 9.53mg/L), 3.04-5.50(inlet concentration of 11.97mg/L, outlet concentration of 0.44-0.26mg/L), and more than 10.72(inlet concentration of 12.24mg/L, outlet concentration of less than 0.47mg/L) under the conditions without precipitator, with Fe2(SO4)3 and with CaCl2 respectively.Introduction Phosphate coating is commonly used for protective coating of machine elements. Including oil removal by caustic washing, rust removal by acid washing, phosphatization and inactivation, the phosphatization process produces wastewater containing inorganic salt of high concentration, especially the phosphate, beyond the environmental standards and polluting environment seriously. What’s more, the wastewater from zincic phosphatizing technology also contains another pollutant, zinc. At present, Chemical Precipitation (CP) and Biological Nutrient Removal (BNR) are two of most common technologies of phosphate removal. BNR, such as A/O, A2/O and UCT technology, is suited for wastewater containing organ phosphorus of low concentration and CP for wastewater with phosphate of high concentration [1]. Researches on CP treating wastewater containing phosphate of high concentration, mainly focus on adding quantity of precipitators[2-5] and coagulants[2-4] and the choice of technologies[3-9]."

Jan 1, 2008

By P. H. Morris

Nineteen field trials of the co-disposal of fine coal tailings and coarse coal wastes were conducted at the dual Goonyella-Riverside mine in Queensland and at the Hunter Valley No. 1 and Warkworth mines in New South Wales, Australia. The profiles of and the sorting that occurred on the co-disposal deltas are described and an existing theory of subaerial deposition, originally developed for mine tailings alone, is shown to apply to the co-disposed wastes. This enables the dimensionless profiles of large deltas, which may take years to deposit, to be predicted from data obtained from a small delta formed in a few days or even hours

Jun 19, 1905

By V. F. Parry

Subbituminous coal or lignite when stored in random riles or in ordinary bins in which the temperature fluctuates and air circulates quickly undergoes objectionable physical and chemical changes, which decrease the utilization value of the fuel and increase its tendency to ignite spontaneously. Slacking takes place, and the heating value decreases. Those changes are accelerated by fluctuations of temperature and by circulation of air through the storage space. When such coals are exposed to the weather they lose moisture rapidly. If evaporation from the surface of a lump proceeds at a faster rate than that at which it is replaced by moisture from the interior, the rate of shrinkage at the surface is greater than that at the interior; consequently stresses are generated at the surface which cause the coal to disintegrate. Likewise, when moisture is absorbed by the air-dried surface the outside layers expand at a faster rate than the interior, which results in further break-down. As these cycles are repeated the lump may disintegrate completely. Therefore, successful storage of low-rank coals depends on control of the factors that influence change of moisture.

Jan 1, 1941

By Stanley C. Suboleski, C. B. Manula

INTRODUCTION In any mining operation all possible steps should be taken to increase efficiency. One area for improvement is mine planning and design, particularly in the area of equipment selection for room-and-pillar systems. Be- cause of the availability of a wide variety of face machines, a fair degree of selectivity can be exercised in the choice of equipment for a particular job. However, this choice must be made on the basis of quantitative facts and forecasts related to the mining application. The purpose of this section is to develop and analyze the details of the mining process. Some specific areas studied include the relationship of system design to productivity, suboptimization as a result of equipment changes, and measurement of system performance. The plan of work leading to a quantitative description of these study areas is based on the growing interest in total system design using simulation as an analytical method (Manula, 1963). CHARACTERISTICS OF PRODUCTION OPERATIONS FROM ROOM-AND-PILLAR SECTIONS For a given mining method, raw production in a given section of a mine is primarily dependent upon the coal seam thickness, roof and floor conditions, methane emission, the mining methods, and the man-machine element. Average section production varies from 300 to 800 st per shift for conventional and continuous mining in high seams and from less than 100 to 300 st per shift in low seams. Since the reject varies from 0 to 40%, these figures must be decreased by the appropriate percentage to reflect the amount of clean coal mined. Personnel requirements per production section per shift for the various methods are listed in Table 1. Table 1. Production Personnel Method No. Method No. Conventional 12-1 5 Longwall 9-14 Continuous 9-1 2 Shortwall 9-12 MINING VARIABLES To evaluate the constraints and interrelationships for various mining methods, it is necessary to categorize the variables which underlie system production potential. Seven critical independent variables which determine production can be identified and categorized (Suboleski, 1978) : Seam Height The five categories are as follows: less than 36 in.;. 36 to 55 in.; 55 to 100 in.; 100 to 180 in.; and greater than 180 in. Floor Quality Floor quality ranges from : Excellent: Smooth, hard, grades less than 1 to 1 % % , and dry. Good: Smooth, soft but dry, with grades less than 3 % . The floor will deteriorate, but cautious operation can prevent it. There may possibly be heaving at some later time. Fair. Soft and damp. There is occasional interference with equipment operation; requires the use of four-wheel drive shuttle cars; ruts with regular use, and may have adverse grades of 5 to 7%. This may be coupled with slippery bottom and/or occasional steep rolls. Poor: Soft and wet. Requires blocking of the bottom to support equipment. There are frequent steep rolls and grades in excess of 7%. Roof Quality Roof quality ranges from : Excellent: Men are able to work under the unsupported top during the initial production cycle if legally permitted. Good: The roof is bolted on a 4 x 4 or 5 x 5 pattern with short bolts (442 in.) or <seam height if the seam >42 in., or requires posting with no bolts on a 4 x 4 or 5 x 5 pattern. There are no falls. Average: The roof is normally bolted on a 4 x 4 or 5 x 5 pattern, but with long bolts (>seam height or >6 ft.). There are infrequent minor falls or there may be an excellent roof which is difficult to drill. Fair: This type often requires spot bolting in addition to the regular pattern or bolting with planks. The roof conditions require shorter than planned cuts, or narrow cuts. Poor: This type requires bolts plus crossbars and posts, or installation of yielding supports or truss-type support. It is almost certain to fall if this is not done. Methane Liberation This ranges from none detected to low (no buildup at the face, even with minimum ventilation requirements) to moderate (the curtains must be extremely tight and tubing close to the face or methane will build up to 1 % during the loading of the car) to high (methane will build up to 170 if the miner is operated at the normal rate, even with proper ventilation). Hardness of Coal Coal hardness falls into the following categories: Soft: Soft coal is easily cut by a continuous miner. A plow could be used by longwall. Average: Coal of average hardness could be easily cut by a miner, and a shearer would be used in the longwall. Moderate: Moderately hard coal causes difficult cut-

Jan 1, 1982

By William A. Beck

The Construction Aggregates Corp. operates a plant for processing sand and gravel from two deposits in Ottawa County, Mich. The bulk of the sand is obtained from a sand dune on the shore of Lake Michigan, and the gravel is dug from a glacial-drift deposit farther inland. Mining is done with draglines, and the material, both at the digging site and in the plant, is moved mainly by systems of conveyors. The Ferrysburg plant serves customers along the Great Lakes, which is unusual for a business normally local in character. Annual production exceeds 1-1/2 million tons of sand and gravel as concrete aggregate and 200,000 tons of core sand for foundry use.

Jan 1, 1962

By D. R. Shoemaker

BACKGROUND AND PROJECT DESCRIPTION A 16.5 kt per day underground molybdenum mine is being developed at Questa in northern New Mexico by Molycorp, Inc., a wholly-owned subsidiary of Union Oil Company of California. Small scale underground mining first began in 1920 and large scale open-pit mining was initiated in 1965, but because of the high stripping ratio, surface mining is being phased out. Construction of the underground mine began in 1979, and first production is scheduled for 1983. Full production of 9 Gg of molybdenum per year will be reached in 1984. Access to the mine will be through a service shaft, development shaft, and a conveyor decline. &apos;The service shaft, 7.3 m in diameter and 396 m deep, will be used for personnel, supplies, development muck, and fresh air intake. It will be equipped with a large, double-deck man and material cage. A double-drum hoist will be used for cage hoisting and will handle up to a 27-t load in balance with a counterweight. Development muck will be loaded into skips in balance and a single-drum auxiliary hoist will provide secondary access. The 396-m-deep, 4.3-m-diameter development shaft will be used for initial mine development, mine exhaust air, and a slick line pass for concrete entering the mine. The development shaft is located 400 m north of the service shaft. Run-of-mine ore will be transported from one of two 1.4-kt storage pockets by a 1.2-m-wide conveyor belt located in the 10°, 1980-m-long, 5.0-m-wide, and 4.6-m-high decline. The decline conveyor belt has a centerline length of 2229 m and a vertical rise of 363 m. It will be powered by four 448 kw motors driving a single drive pulley. There will be two 1.2 m transfer belts on the surface with centerline lengths of 322 m and 542 m. All three conveyor belts will be reinforced with steel cables. On the surface, in addition to the headframes and hoist houses for the two vertical shafts, a large maintenance shop, warehouse, compressor house, change house, main office building and timber framing building have been constructed.

Jan 1, 1981

By M. Massinaei, M. R. Yarahmadi, S. Zeidabadi, M. Oliazadeh

"Although the majority of new flotation column installations continue to be for cleaning or recleaning, an increasing number of flotation projects are specifying columns for roughing or scavenging duties. NICICO has recently installed flotation columns operating in rougher, cleaner and re-cleaner circuits at three new copper concentrators; Miduk, Sungun and Sarcheshmeh. Miduk is probably one of the unique operations that use flotation columns in a rougher circuit. The Miduk rougher circuit comprises five CISA parallel columns. The rougher concentrate constitutes 88% of final concentrate and reports to concentrate thickener. The rougher tailings are scavenged with ten 50 m3 mechanical cells. The scavenger concentrate is then reground and cleaned in three 3.2 m columns in parallel. The cleaner tailings are scavenged in six 50 m37mechanical cells. The cleaner tailings along with the rougher scavenger constitute final tailings.In Sarcheshmeh, the cleaner concentrate is re-cleaned in a column. The column tailings are sent back to the cleaning circuit and the concentrate is reported to the concentrate thickener.This paper reviews performance details of the flotation columns utilized in rougher circuits at Miduk and compares their metallurgical performances in an industrial scale with columns at cleaning circuits for copper flotation. Practical limitations for applying the flotation columns in rougher circuits will be discussed in details.INTRODUCTIONThe use of flotation columns has grown rapidly since the first successful commercial application at Mines Gaspe, where columns were used in a cleaning circuit to float molybdenite. Early applications of columns were restricted to replacing or adding final cleaning stages in existing circuits. Columns can achieve superior cleaning as compared to conventional cells due to their ability to support a large froth depth and the use of water for washing froth. Wash water travels down through the froth and crosses the pulp/froth interface, replaces the entrained particles reporting to the froth from pulp and improve the concentrate grade."

Jan 1, 2011

[Page Minutes 7 Discussion of L. S. Ickis&apos; Paper 20 City Planning and Landscape Architecture for Western Coal Camps, by S. R. DeBoer 22 Discussion of Thos. Foster&apos;s Paper 26 Discussion of Safety Committee Report 30 Discussion of R. L. Hair&apos;s Paper 52 Discussion of H. I. Smith&apos;s Paper 61 Miscellaneous Discussion 62 Discussion of Thos. A. Stroup&apos;s Paper 70]

Jan 1, 1924

By S. Hardy

The paper describes state of the art non-linear analyses of the pile groups that were employed during the raising of the 133m high triumphal arch which forms the centre piece of the new Wembley stadium in North London. The deformation behaviour of these piles groups was critical during the lifting process, due to the slendemess of the arch structure. The issue was made more critical by the high profile nature of the project. The results of Class A (Lambe (1973)) predictions of the pile group movements are compared with observed displacements and rotations which were accurately measured on site. To illustrate the importance of soil non-linearity on pile group behaviour, comparison is also made with pile group analyses which assume a linear soil stiffness.

Jan 1, 2006

By M. Y. Solar, W. G. Davenport

Most of the recent growth in nickel demand has been for the production of stainless steels and virtually all of it has been satisfied by new ferronickel and nickel pig iron projects. But "good" laterite resources are more and more difficult to find: Ni grades are decreasing while Fe contents are increasing and SiO2/MgO ratios vary from 1 to 3 and even higher if limonites are included. Predicting the smelting behaviour of such ores requires a more fundamental understanding of the thermodynamics of these systems in commercial operations. One valid question is thus whether these furnaces are at thermodynamic equilibrium or not. The present paper investigates this question, compares calculated and measured O2 partial pressures and mechanical/chemical slag loss ratios and concludes that ferronickel furnaces are not at thermodynamic equilibrium. This departure from expectation is actually to the operators’ advantage: the targeted ferronickel grades can be obtained at lower reduction potentials than calculated from pure thermodynamics.

Jan 1, 2014

Validation and Simulation of Optimised Compressive Crushing Ores and aggregates are used in countless applications all around in the society. The frequent and escalating usage of these natural resources requires efforts aimed at not only making do with what is available, but also undertaking actions to improve the production yield and the processes involved. This strive to improve the performance and efficiency of existing crushers and crushing processes has brought about research concerning optimisation of crushing plants as well as theoretical optimisation of crushing. Previous work shows that different applications and different materials require different compressive crushing in order to accomplish a theoretically optimal size distribution. These optimisation results, obtained by genetic evolutionary algorithms and evaluated with so called fitness functions, convey the theoretically optimal number of compressions and their corresponding compression ratios. Presented in this paper are thus some laboratory crushing tests, which are conducted in order to validate a set of previously optimised crushing sequences. More specifically, interparticle crushing sequences are carried out in a Piston-die testing equipment, where samples are sieved between each compression. The subsequent analysis of these results shows good correlation with model predictions, which combined with a comparative study between presented optimisation results and measured data from a real crusher indicate a significant potential for improvement. However, before redesigning crushers according to what is theoretically optimal, the resulting impact as well as the effects on the outcome of a crushing plant must be examined. Crushing plant simulations using the software PlantDesigner« are consequently performed. Models of existing crushers are altered, resulting in changed particle size distributions from some of the crushers in a crushing plant. Thence, it is concluded that significant gains can be made in terms of capacity, process yield and energy consumption.

Sep 13, 2010

By V. Arslan

The primary advantage of using oxalic acid is in the formation during leaching of soluble complexes that are both photo-chemically and microbiologically degradable. In this study, chemical leaching and bioleaching methods were used for quartz sand. Using the ANOVA-Yates test technique, the effects of solids rate, temperature and oxalic acid rate were investigated using chemical leaching experiments. Solids rate and temperature were effective parameters on the chemical leaching. After the tests, the removal of Fe2O3 increased to 97.02%. Finally, the iron content decreased from 0.315% Fe2O3 to 0.0094% Fe2O3 in the leached quartz sand. The effect of pulp density was investigated on the bioleaching experiments. The bioleaching of iron oxides by Aspergillus niger resulted in removal of 47.70% of the total iron in the quartz sand, and suggested that this strain has not a good ability to remove iron impurities from quartz sand in comparison with chemical leaching. The iron content decreased from 0.315% Fe2O3 to 0.164% Fe2O3 in the bioleached quartz sand.

Jan 1, 2009

By David Colloby, Darelle T. Janse van Rensburg

Premature coating failures of new and refurbished structural steelwork of clarifiers have been a common occurrence at power generation plant in recent years. Several of these failures have ended in disputes regarding which party was actually to blame for the early degradation of the coating system, i.e. quality of the power station?s water, poor application, substandard materials, outdated corrosion protection specifications, etc. Having investigated several of these premature failures, it is clear that the corrosion protection of clarifier steelwork is a highly complicated process and that the painting work needs to be closely monitored in order to ensure long-term protection by whichever product that is applied. The need for quality control inspections during the painting of the plant is often underrated during the project planning stage, also the importance of detailed corrosion protection specifications that consider not only the practical application of the materials, but also the quality control responsibilities of the Contractor, End User and Third Party Inspectorates that may be involved. This paper considers the application constraints of coatings to clarifier structural steelwork and importance of stringent quality control procedures in order to ensure the long-term performance of the corrosion protection systems.

Jan 1, 2006

By Tapiwanashe Chenje, David Hewitt, Peter Radziszewski

"Steel media wear is a major cost component in mineral processing. In the case of existing operating mill, steel media wear prediction can be used to help compare the performance of steel media from different suppliers before completing marked ball tests. It can also be used to investigate how ores from different site of a given mine affect steel media wear. In the case of Greenfield cases, steel media wear prediction can be used to estimate the cost of steel media for a given feasibility study.This paper presents a wear testing and prediction methodology that ties together abrasion wear and corrosion wear testing results through the use of a simplified discrete element charge motion model. The experimentation dimension to this methodology takes samples for steel media, mine site ore and water to determine abrasive and corrosive wear parameters of a given media, ore, water set.The test methodology development is illustrated along with a comparison of wear predictions with observed wear rates in some operating mills. A case study is included that illustrates how this methodology was used to estimate steel media wear performance. The paper closes with a discussion on the use of parts of the experimental methodology to investigate other abrasive wear conditions such as liner wear, paste backfill piping and ore breakage as well as avenues for future test development.INTRODUCTIONThe cost of comminution processes represents a significant part of the overall costs of a mining operation. Grinding media/liner wear alone accounts for 15-25% of the total operating cost. The annual steel grinding media consumption is approximately 2.5 mtpa or 0.3% of total world steel production [Smorgon Steel Group, 2005]. Using an estimated grinding media cost of about US$0.75/kg [Halbe and Smolik, 2002], gives a media cost of over US$1.875 billion per annum. Mining continues to shift towards exploitation of lower grade ores as the higher grades get depleted. These low-grade ores require finer grinding for the liberation of the value minerals, and hence it can be assumed that grinding media consumption will also continue to rise.The Bond abrasion test, dating back to 1963, [Weiss, 1985] is still used to estimate the total mill media wear. This is a purely empirical test based on the ore. From Bond abrasion index tests, it is possible to determine a wear rate using the following function:"

Jan 1, 2009

By Art Cooper

Granulation systems are common in the metals industry. The process of granulating slags, mattes and metals is more than a century old, and a number of different systems have been developed to accomplish this task. These systems have followed an evolutionary path, and may be characterized by common elements in terms of molten material handling, granulation methods and post-granulation processing. This paper will summarize the more common systems used in the metals industry, with particular reference to non-ferrous applications.

Jan 1, 2006

Terramin Australia LimitedÆs Angas Lead and Zinc Mine operates in the unique and picturesque environment of Adelaide Hills, 60 km South East of the Adelaide city centre itself. The mine was granted approval for construction by the South Australian Government in April 2007 and the Process Plant was constructed and commissioned in July 2008. This paper outlines the background to the plant design and the approval process given the close proximity to the local town of Strathalbyn, with a major focus on the commissioning learnings from the Concentrator startup for future reference and a review of the operational performance of the Concentrator throughout the first year of successful production versus design and expectations.

Jan 1, 2009

By Gilles Landry, Mohamed Ourriban, Pierre Pelletier, Ahmed Bouajila, Claude Gagnon

"This paper summarizes the work carried out to develop and implement the gold cyanidation by size class (CSC) process, a novel concept aiming to reduce cyanide consumption.First, the authors will present the methodology and results of laboratory tests conducted to evaluate the response and the expected gains for the selected ore when submitted to the CSC processing. In the second part, focus will be on the methodology behind the CSC scheme development prior to its implementation at Sleeping Giant Mine concentrator. Finally, the evaluation of the modified circuit performance (operating since July 2007) will be addressed as a reduction of cyanide consumption by more than 30% and of free NaCN in the tailings by more than 50% is confirmed.INTRODUCTIONCyanidation is the most widely used technique to extract gold from hard rock ore. Its use of oxygen as an oxidant still makes it the cheapest process available after more than a hundred years. Unfortunately, gold cyanidation is known to be a slow process that requires high consumption of cyanide. In fact, cyanidation plants are designed with very long residence times (Habashi, F., 1967). Drawbacks of long cyanidation times are, among others, loss of gold in tailings, high reagent consumption and high capital investment and maintenance costs.Responding to their members expectations for the improvement of cyanidation process, COREM undertook a challenging R&D program covering the development of alternative concepts (processes) and the understanding and solving of mass transfer, gold passivation and process water chemistry problems. One of the interesting findings of the R&D program is a patent pending process technology based on the cyanidation by size class (CSC) concept (Côté and Ourriban, 2004; Gagnon and Ourriban, 2007; Gagnon and Ourriban, 2004).The new technology was tested at the Sleeping Giant concentrator, a plant member of COREM. The existing flowsheet was modified in order to implement the CSC process in the full-scale plant. The plant has been operated using the modified CSC based process since August 2007. Laboratory testwork was conducted prior to the industrial application in order to properly design the new process scheme. The modifications included:1) Reclassification of the grinding circuit product through an hydrocyclone2) Leaching of the coarse size fraction3) Addition of the fine fraction to the coarse leaching circuit tails."

Jan 1, 2009

By G. V. R. Landman

It is self-evident that economic pressure demands that today?s mines stay ahead and stay competitive. The challenge, however, isto find fresh approaches to the pursuit of efficiency and operational effectiveness. Rather than focusing on individual activities in isolation, a process that can lead to poor investment decisions affecting other critical areas of operation, a key concept is optimization of the whole value-adding process of mining. This concept is often called ?mine to mill?. Whereas it is never easy to establish optimal conditions across a whole mining operation, the quality and consistency of the blasted rock strongly affects all operations. In many instances an inability to control rock breaking has been unchallenged, leading to significant investment in time and money to manage the consequences of poor and/or inconsistent rock breaking. This paper discusses how electronic initiation systems can play a major role in optimizing the whole mining operation. Electronically programmed detonators provide not only the ability to tightly control a key parameter, namely timing, but also the opportunity to extract blast data for a continuous improvement process. The ability to alter predictably, using feedback from the downstream processes, blasting variables such as fragmentation, throw, wall stability, and vibration, is a huge advantage. The current understanding of the effect of timing, accuracy, and scatter on rock fragmentation is reviewed and offers reasons why there is significant progress in controlling blast results. Two case studies are examined, which demonstrate the benefits of the technology on the economics of the operations under review.

Jan 1, 2010

The area around Big Bell was actively explored and sporadically mined for gold from 1904 onwards. The lease was named Big Bell in 1913. The property became established as a major mining centre with the development of the Big Bell underground mine in 1936. Mining operations commenced in late 1937 and ceased in January 1955, after milling 5.56 million tonnes of ore and producing nearly 21 tonnes of gold. In 1969 Australian Consolidated Minerals Ltd (A.C.M.) acquired the deposit then explored it alone and in joint venture until the current joint venture agreement with a subsidiary of Placer Pacific Limited (PPL) was negotiated in 1984. Following exploration and evaluation of near surface mineralisation, and a reassessment of the remnant pillars underground, the concept of a large open pit in conjunction with an underground mine evolved. The feasibility study was completed in October 1987, and the decision to go ahead was formally announced on 2nd December 1987. Placer Dome Technical Services Ltd were contracted to be the Project Managers, commencing construction on site early in January 1988 and handing the plant over to the operators 13 months later. The capital cost of the project at handover was estimated to be $133 million. The operating team commenced pre-production stripping of the open pit on 1st May 1988. By 31st January 1989 almost 16.0 million tonnes had been stripped and 350,000 tonnes of ore stockpiled.

Jan 1, 1989

By H. E. Lee, D. Ingvoldstad

AT Bunker Hill the original charge storage and preparation system was installed in 1917 to accommodate lead-silver gravity mill products. Only minor tonnages of wet fines such as vanner and flotation concentrates were received. Charge flux and diluent requirements were provided by an ample supply of siderite middlings and coarse lime-rock. While oversize material was crushed through %-in., the use of roll crushers in series resulted in a more or less granular, free-running charge of adequate porosity. Under such conditions, receipts could be proportioned directly from receiving bins, and suitable blending was obtained by passing the resultant layered composite through a small Stedman disintegrator. This elementary system served well for production and smelting of an extremely lead-rich sinter, in the low column blast furnace. Here sinter physical quality is less critical, and appreciable quantities of raw flux and oxidized ore may be charged directly to the blast furnace without seriously impairing furnace capacity. Prior to 1938 the relative proportions of zinc receipts were low and extraneous flux requirements at a minimum. However, subsequent war-inflated zinc demands brought marked increase in fine concentrate and slimy zinc leach residue receipts and an abrupt introduction to smelting a more refractory zinciferous charge. The higher temperature demands of the more refractory zinciferous charge required a higher smelting column and the higher column, in turn, rendered the blast furnace less receptive to improperly prepared feed. Inadequate presinter processing facilities prevented the production of acceptable sinter, and blast furnace production declined to 50 pct of former capacity. In view of a prevailing contention among lead metallurgists as to the deleterious action of high zinc slag on blast furnace capacity, it was natural to overstress this factor and to concentrate on possible changes in furnace design. Another interpretation was that primary correction should be sought through expanded sintering facilities. This divergence of opinion, together with the extreme material shortages of the 1940&apos;s, delayed corrective

Jan 1, 1957