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Decrease dust concentra¬tions of longwall face workers by keeping them up¬wind of the headgate stage loader and crusher. Approach Direct the face airflow from tailgate-to-headgate, the direction of coal transport. How It Works Ventilation is the primary method of dust control on longwall faces, and the most effective ventilation control is to put workers upwind of dust sources. Recent ad¬vances in longwall dust sampling by the Bureau have shown that on some faces, the headgate crusher and transfer points are signifi¬cant sources of dust -some¬times making up more than half of the total dust to which the face workers are exposed. Also, when the air is coursed from head-to-tail, the relative velocity between the air and moving coal on the face conveyor can cause dust entrainment from the conveyor, particularly if the coal is dry. On the faces where these dust sources are a signifi¬cant problem, they may be eliminated by coursing the air from tailgate-to-headgate, provided that the mine roof is stable enough to permit the necessary intake of the tailgate along with an aux¬iliary intake at the headgate.

Jan 1, 1982

By I. L. Nichols

Laboratory characterization studies have investigated the processing variability of carbonaceous uranium-bearing materials found in-sandstone deposits. This work was conducted by the Federal Bureau of Mines in support of its goal to maintain an adequate supply of minerals to meet national economic and strategic needs. Conventional processing produces only partial uranium recovery from these ore materials. Ore samples from New Mexico, Texas, and Wyoming were examined, and each responded differently. Using direct acid leaching techniques, some of the ores required acid additions of only 50 pounds per ton, while others required 300 pounds per ton of material. Maximum extractions ranged from 83 to 94 percent in the six ores tested. Roasting those ores prior to leaching improved extraction of the uranium to the 90 to 99 percent range. An oxidative autoclave treatment used with five of these ores resulted in extraction of 78 to 97 percent of the uranium. The experimental studies also showed that some carbonaceous materials can be concentrated by froth flotation. The flotation tailings were amenable to acid leaching, and up to 98 percent of the uranium was extracted from this type of material.

Jan 1, 1979

By Warren Blauvelt

IN VIEW of the enormous wastes of natural resources, of labor and of capital, due primarily to the economic environment, established by legislation, the general neglect of this phase of the problem of the elimination of waste is difficult to understand. In such a study, the engineering method of attacking a problem should be employed. The engineer does not try to solve his problem by disregarding natural laws, but by utilizing them. The successful engineer establishes an environ-ment in which forces, acting in harmony with natural laws upon an object or objects, will produce the desired results. The engineer who notices that the water level in a vessel, into which a gallon of water per minute is flowing, is two feet higher than the level in an adjacent vessel, connected therewith by a 6-in. pipe, does not assume that the law of gravity has ceased to work, or that water no longer seeks its level; he examines the connecting pipe for obstacles which prevent the free flow of water. Some editors, many reformers and most politicians tell us that the law of supply and demand no longer works; hence governmental control of commerce is necessary. Governmental control may temporarily-that is, until trial proves its futility- "insure domestic tranquility" and "promote the general welfare;" but it cannot "secure the blessings of liberty" which it cur-tails, nor can it "establish justice" which it violates.

Jan 9, 1922

By Sadan Kelebek, S. O. Fekete, Gerhard W. Heinrich, Peter F. Wells

"For many years, rejection of pyrrhotite at the milling stage has been a viable method for sulphur dioxide abatement. Pyrrhotite rejection options have included the use of selective pyrrhotite depressants in flotation; a depressant which has recently come into use is diethylenetriamine (DETA). This reagent has been the subject of recent research and development efforts in the processing of nickel-copper sulphide ores.In the present study, adsorption trends for DETA have been monitored under the dynamic conditions of flotation using a spectrophotometric method which has been adopted for this purpose at Falconbridge Limited. The results indicate that the adsorption of DETA is quite fast and in the majority of cases, accounts for the disappearance of 75-95 % of this reagent from pulp's liquid phase. The adsorption decreases linearly with an increase in the DET A dosage from 100 git to 600 git. The tailings produced in flotation tests were subjected to washing with deionized water at 60 °C to examine the desorption of DET A. These results indicate that the amount desorbed increases with increasing DETA dosage used during flotation but is less than 7 per cent of that adsorbed. It was found that adsorption of DETA alone, even at relatively high levels, does not always cause depression of pyrrhotite. A new combination of reagents identified in the metallurgical laboratories of Falconbridge Limited utilizes sulphur dioxide with DETA. These reagents together are capable of efficient depression of pyrrhotite on a consistent basis."

Jan 1, 1995

By Dundon P

The Broken Hill Proprietary Company Limited (BHP) is Australia's largest company with widespread commercial interests ranging from steel productin to minerals, oil and gas recovery. While these interests are worldwide, the Company operates or is partner in more than 30 Australian mining ventures. BHP Engineering (BHPE) is the in-house consultant to BHP but is also the largest consultancy in Australia available to general industry. Australian Groundwater Consultants Pty Limited (AGC) is a large, independent consultancy offering specialist services in all aspects of water resources, waste management and environmental engineering, both in Australia and overseas. AGC is particularly well regarded within the mining industry and has a considerable record of providing solutions for mine water related issues. By drawing on the extensive combined experience of AGC and BHP Engineering, this paper will summarize some of the unique ways in which water interacts with and impacts on mining operations in Australia, particularly in matters such as: Slope Stability Drainage and Dewatering Production Cost (Transport, Drilling, Blasting, etc) Water Supply Water Management Tailings and Waste Management Environmental Issues. The paper will also illustrate the cost benefit of providing major engineering works for reduction in total mining cost. It will be shown that, through advance planning and close interaction between outside experts (ie consultants) and on- site personnel, very significant operational cost savings can be made while at the same time achieving a safe and responsible mining environment.

Jan 1, 1988

By Fisher WJ

The'treatment of the South Alligator valley uranium ores passed through three phases dictated by economic circumstances during the years 1956- 640 In the first phase high grade ores were treated by hand sorting and wet gravity separation for recovery of high grade concentrate to meet a United States Atomic Energy Commission (U.S.A.E.C.) contract for plus 50 per cent pitchblende concentrates. Secondly tailings from the gravity plant and some high grade ores were trucked to Rum Jungle and sold to the Australian Atomic Energy Comm- ission (A.A.E.C.). Thirdly solvent extraction plants were built at Moline (previously Northern Hercules) by United Uranium N.L. (UOU.N.L.) and at the South Alligator River by South Alligator Uranium N.L. (SOA.U.N.L.). These were the first solvent extraction plants designed in Australia, as earlier plants used the ion exchange process for recovery of uranium oxide. The plants were small, indeed the S.A.U.N.L. 15 ton per day plant was possibly the smallest commercial plant known, but they operated efficiently with good recoveries, and were placed on a care and maintenance'òbasis when contracts were filled. Sulphuric acid was supplied in bulk to both plants by the A.A.E.C. from the 30 ton a day acid plant at Rum Jungle. Gold was recovered as a by product of uranium ore from the Moline plant.

Jan 1, 1968

By Henk Dijkman

Existing metallurgical operations can be optimized by implementing engineered bioreactor systems in the process configuration. For instance, metals can be recovered safely and economically from process and waste streams using biogenic sulfide. This paper focuses upon possible applications of high rate biotechnology in metallurgy and mining. Metal recovery from liquid streams (process bleed streams, leach water, waste streams), integrated removal of (fugitive) SO2, treatment of acid blowdown and other options will be described. Practical examples are given. A large-scale application of biotechnology at Pasminco's Budel Zink refinery is described. Here, acid plant blowdown is treated together with electrolyte magnesium bleed. Concentrated zinc sulfide (10 tpd) is produced biologically, eliminating the production of gypsum and closing the refinery sulfur cycle. Recovery of metals from solution at Kovehute Pribram in the Czech Republic is the second practical example. In this case, H2S is produced on-site for metal removal from an alkaline slag dump leachate. At the Caribou mine in New Brunswick, Canada, acid mine drainage is treated with biogenic sulfide. This plant is situated upstream of a lime neutralization plant. As a result, reusable ZnS is produced, the volume of the sludge (gypsum and hydroxides) from the lime plant is reduced by 50%, the sludge is free of heavy metals and the final effluent concentration of heavy metals is in the ppb range. In the second phase of this project, metals will be recovered from stored tailings, resulting in significant revenues.

Jan 1, 2002

By R L. Bowers, D S. Smit

In light of soaring commodity prices, Nkomati has investigated the potential of processing additional low grade nickel ores to add to their reserves. The Chromititic Peridotite Mineralised Zone (PCMZ) was identified as one ore type with the potential of producing a high grade nickel concentrate. In addition to the nickel, copper, cobalt and PGM?s, the ore also contains a significant amount of chromite which adds substantial value to the project. Pyrrhotite, pentlandite, chalcopyrite and chromite are the dominant minerals within the ore, which on average grades around 0.23% Ni and 12% Cr2O3. The PCMZ ore contains excessive amounts of fibrous silicates which makes it a challenging float, with gangue control as the key element to success. Extensive laboratory testwork, which eventually led to a pilot plant campaign, was undertaken with the aim of developing a process route for treating the PCMZ ore. Several flotation circuit configurations were tested and results from the optimal circuit indicated that a nickel concentrate grading at 8.5% is achievable at a recovery of 78%, with substantial recoveries for the copper, cobalt and PGM co-products. The chromite concentrate, grading at 45%, was extracted from the flotation rougher tailings via a spiral plant. These promising results were incorporated into the plant design and are currently being evaluated in a bankable feasibility study, which will see the current 100ktpm MMZ concentrator being upgraded to a 250ktpm PCMZ concentrator as part of Nkomati?s Phase II expansion.

Jan 1, 2007

By N. Katsikaros, R. Agius, B. Adaszynski

"With stricter limits globally being imposed on cyanide discharge levels in plant tailings, having a reliable and well designed cyanide detoxification circuit is critical for any operation. The INCO SO2/Air method is most commonly used for cyanide detoxification where sulfur dioxide and oxygen combine to oxidize cyanide to the much less toxic cyanate. This paper discusses the steps involved with the development, from laboratory through to design, of a robust cyanide detoxification circuit. The goal is to produce a cyanide detox circuit that reliably meets discharge target with the lowest operating and capital cost.Design begins in the laboratory where continuous tests simulate full scale operation of a detoxification circuit. Detailed design relies on several design parameters, based on theory and experience, to ensure the reactor can satisfy requirements of the cyanide destruction reaction. Selection of the correct reagents and their supply to the reactor is important to an operable process with minimal operator workload. The operability of the system relies on correct process control strategies to maintain optimal reagent addition and chemical conditions in the reactor. These design factors are critical to providing a robust plant that will consistently meet cyanide discharge levels, ensuring uninterrupted production and plant profitability."

Jan 1, 2014

By G. Lodewijks

Today most belt conveyors used for the transportation of bulk solid materials are driven in a centralized way with all drives in one location. The dynamics of these systems has been studied since the early seventies of the previous century and are well understood. Compared to the more conventional belt conveyor system with a centralized drive station positioned at the head and/or tail of the system, a decentralized driven layout with drive stations distributed along the whole length of the belt offers a number of advantages. However, to make a decentralized driven configuration a good alternative to a centralized driven, the system designer will have to balance these benefits with the increased complexity. The full potential of a decentralized driven system can only be realized if the right balance is found between the locally applied drive force and the motion resistances occurring along the system. From previous research it was found that in a multiple driven belt conveyor system the most dominant belt behavior occurs in the belt section between the tensioning device and the first drive station. In this paper different possibilities are analyzed to improve this dominant belt response and optimize the starting behavior. This can be accomplished by changing the predescribed start up profile, starting the drive stations in sequence rather then all at the same time and altering the locations of the drive stations.

Jan 1, 2008

By R. G. Mallon, A. E. Lewis, R. L. Braun, C. J. Sisemore

Large ore-bodies, too deep to mine economically by underground methods, may be mined using nuclear explosives and a high pressure leaching technology. Breaking of a hypothetical ore-body with a series of nuclear explosives is described. The leaching would be accomplished by pumping oxygen into the flooded rubble under the normal hydrostatic pressure. The injection of oxygen gas into the bottom of the rubble chimney induces a circulation of the solutions which carries dissolved oxygen to all parts of the chimney. Copper and other metals would be recovered from the solutions by solvent extraction as they are circulated through a surface plant in a closed system. This provides for control of pH as well as containment of radioactivity. Radioactivity levels in the solution are low, and no radioactive impurities are expected in the copper. This technology will allow production from deposits too deep to be of interest with present technology and will reduce or eliminate some environmental effects of conventional mining, such as open pits, mine dumps, tailings ponds, and SO2 effluent. Close proximity of cities and towns is a limitation because of seismic damage. The technology requires further development for commercial application, and exploration targets must be enlarged to include deeper deposits, before the technology can be applied on a large scale.

Jan 1, 1974

By H. B. Treweek, V. R. Degner

Current mineral industry trends toward handling the high tonnages of lower grade ore and tailings retreat lead to a continuing need for large volume flotation cells. The benefits which accompany the larger cell flotation circuit include reduced capital cost, less mill floor space, reduced operational complexity, and less maintenance, provided the metallurgical performance and solids suspension capability of the cell are not compromised. This study reviews development considerations pertinent to large mechanical air induction type flotation cells from the standpoint of cell internal hydrodynamic characteristics and metallurgical performance. While the importance of process related factors such as chemical usage, pH control, solids size and loading, and circuit arrangement cannot be denied, this study will emphasize the relationship of cell size (volume), mechanism geometry (impeller diameter), and mechanism operating conditions (rotor speed and submergence), to air transfer, liquid circulation, liquid "rise velocity" and mechanism power draw. Specific hydraulic performance maps for large capacity 300 and 500 cu ft flotation cells will be reviewed in terms of the relationship between the operational variables and the air and liquid flow conditions within the cell. Metallurgical results, obtained under plant operating conditions for 300 and 500 cu ft cells, are included to illustrate the metallurgical performance similarity of flotation cells which follow hydrodynamic scale-up criteria set forth in this investigation.

Jan 1, 1976

By Hinkley JT, Smitham JB, Nicol SK

Historical investigations of mineral flotation have highlighted the many factors which affect flotation. Some of these are related to the mineral, its degree of liberation (however defined), solution chemistry and flotation reagents. The complexity of the interactions between these factors often leads to uncertainty over whether the final grade/recovery results obtained in either laboratory or pilot scale tests are in fact the maximum recovery at a specific grade. The reason for this is that there has not previously been a general technique for establishing the floatability of an ore. This study describes the further development of the 'tree' flotation technique for minerals other than coal, and in particular its application to iron ores. Oxides have been chosen as a system of intermediate complexity between the simple coal system and sulphide minerals to illustrate the principles and develop the concepts. The 'tree' flotation procedure consists of the progressive flotation to completion of successive froth products, followed by reflotation to remove entrained material. It therefore represents an 'equilibrium' approach to floatability rather than a kinetic approach. By ordering the various products and tails in order of increasing grade, a cumulative curve of grade and recovery can be constructed against which an isolated batch or plant result on the same ore can be compared. In this investigation, a Jaspilitic iron ore was tested at various grind sizes to determine the 'best' flotation grade-recovery that could be obtained with a given set of reagent conditions. The separation characteristics of a process are best measured on a system which contains particles which behave as 'middlings' with respect to the process in question. For this reason a natural ore was chosen in preference to model systems in order to relate more readily to plant situations.

Jan 1, 1993

By Chris J. Martin, Nichola A. McKay

"In mid-2001, North American Palladium commissioned a new 15,000 tonnes per day milling and flotation circuit at their Lac des Iles Mine (LDI), one of only two primary platinum group element (PGE) producers in North America. The mill treats an ore containing 2.3 g/t PGE producing a concentrate assaying 300 g/t PGE, at a recovery of roughly 75%.Top-size control crushing is conducted ahead of semi-autogenous milling. A portion of the SABC circuit feed is crushed to –25mm to increase mill throughput. Product from the SABC circuit is fed to flotation. Flotation consists of roughing and scavenging, with the rougher and scavenger concentrates each being reground using vertimills. Cleaner flotation uses two stages of cleaning using mechanical cells, and a third stage using flotation columns, the column tails being scavenged using a bank of mechanical cells. Rougher concentrate by-passes the first cleaner circuit. Reagents added include amyl xanthate collector, dithiophosphate promoter, MIBC frother and a carboxy methyl cellulose-based gangue depressant.SGS Lakefield Research, together with staff at LDI initiated a structured plant optimisation program comprising on-site testing, bench-scale flotation and mineralogical analysis using QemSCAN. The objectives of the mineralogical studies were 1) to identify feed PGM, their size distribution and mineral associations, 2) establish relationships between the PGM mineralogy and PGM floatability, and 3) characterize the culprit silicate minerals diluting the final concentrate.PGM flotation response at Lac des Iles is clearly bi-modal in nature, with about 60% of the PGM floating quickly, to a high value rougher concentrate. The mineralogical work was able to explain the causes of the bi-modal PGM floatability, which in turn created new opportunities for the metallurgists to optimise the circuit."

Jan 1, 2003

By H. W. Hitzrot

THE new 4800-ton fine-grinding plant at the Hollinger Consolidated Gold Mines property at Timmins, in the Porcupine district of northern Ontario, went into operation in November 1937, and represents the culmination of an experimental program. The plant is a moderately low rectangular building of hollow tile on a steel frame, large enough to house four 7 by 15-ft. ball mills, with a 12 by 27-ft. classifier beside each mill. At one end is space for the addi-tion of another mill and classifier or a mill and classifier for the regrinding of table concentrates. This regrinding is now done in the old grinding department of the main mill building. Also within the new building are the four 6-in. sand pumps for lifting the product from the low-dis-charge ball mills up to the classifier feed launders and the containers and appurtenances of a complete dust-collecting system. Together with the usual overhead crane, these constitute the major units of equipment within the new plant. The location is at the lower end of the main mill buildings toward the old tailings pond, now so neatly leveled off and sodded to rival the very best of our own W.P.A. projects. Close by the new grinding plant is a large dome-shaped structure-the new fine-ore bin designed and built by Mr. Longmore and the Hollinger milling and engineering staffs. There are four separate automatically loaded inclined belt conveyors leading up from the bottom of the bin to the grinding plant. Each belt is equipped with a recording weightometer, in the grinding building. The whole structure and the machinery within it give the impression of permanence and a carefully integrated design with close attention to details, giving the operators every advantage for consistently good work.

Jan 1, 1939

By J. D. Mooney, D. L. Darnell

The accompanying chart has been drawn for the convenicncc of engineers as' a means of quickly determining the correct number of plies of conveyor belts operating under specific conditions. The calculations are based on the average safe strength (factor of safety, 15) of the various standard rubber conveyor belts. The calculations assume maximum loading conditions; that is, the belt is considered as carrying the greatest load that it will handle without spillage at ordinary belt speeds. This not only produces the most economical operating conditions, but also the maximum tension in the belt. The chart is a graphical representation of the formula: Where, p = kg W(L + 10H) p = the correct number of plies k = a constant, depending on the type of drive g = the weight 'in pounds per cubic foot of material handled W = the width of the belt in inches L = the length of the belt in feet (approximately twice the center distance). H = the difference in elevation between the head and tail pulleys, in feet. For a simple drive, with a bare pulley, k = ----1 For a simple drive, with a rubber-lagged pulley, k = --- 1 For a, tandem drivej with bare pulleys, k = 1/375.000 For a tandem drive, with rubber-lagged pulleys, k = ---1 The chart is drawn for a simple drive, with a bare pulley (k — =e)1 ; therefore, the number of plies obtained from the chart should be multi- ' plied by the factor 0.83 or 5/6 for simple, lagged drive; the factor 0.67 or 2/3 for tandem, bare; and the factor 0.55 or 11/20, for tandem, lagged. The formula p = kgW(L + 10H) has been developed mathematically

Jan 1, 1916

By Hung Tran, Denis Fleury, (Ken) Qian Liu

Pastefill technology has gained an increasing popularity over the past decade. The main advantages are: environmental performance (mainly in tailings disposal), low operation costs and ease of operation, as well as homogeneous properties. However, its weaker nature compared to other types of fill makes it more vulnerable to blast damage. In the last few years, Louvicourt Mine has experienced some cases of pastefill sloughage during the mining of secondary stopes. In order to reduce pastefill sloughage, some studies have been undertaken to investigate the causes and to find possible solutions by looking at stope design, extraction sequence, ground support and blast design. This paper presents the results of a blast damage study carried out at Louvicourt on the 860m level. The instrumentation included vibration monitors (using 6 sets of triaxial geophones grouted in the pastefill), a borehole camera to observe the damage in the pastefill, and a cavity monitoring system for three-dimensional surveys of the blasted stope. The test stope was mined out by four slot blasts and three mass blasts. Most of the damage was created by the mass blasts. It was found that the accumulation of ground vibration was mainly due to the low seismic velocities in the pastefill and the relatively tight delay between blastholes. In addition, there was a substantial difference between the critical PPV in the confined and unconfined pastefill. In practice, these findings have led to modifications of blast design to minimise the exposure time of pastefill walls to air. Such modifications included changes of blast sequence in secondary stopes and delay layout between heavily loaded blastholes. Over the paste two years, the implementation of new blast sequence has shown encouraging results in terms of reduced pastefill sloughage and dilution.

Jan 1, 2000

ARTHUR CROWFOOT, Morenci, Ariz. (communication to the Secretary*).-In order to set at rest any doubts in regard to the reasons for selecting the Callow pneumatic flotation process for use in the concentrators of the Arizona Copper Co., Ltd., at Clifton and Morenci, I am presenting the following facts gained from the results of test work carried out on the Callow and Cole-Bergman types of flotation machines at Morenci. Prior to the comparative test of merit between the Callow and Cole-Bergman machines, many months' time and a large amount of money had been expended in the development and testing of the Cole-Bergman machine, and it was necessary to decide quickly what make of machine was best to install. The Callow pneumatic flotation process was recommended for installation in our concentrators for the following reasons: The results indicated: 1. That the Callow system made the highest extraction of copper values from the feed treated. 2. That the Callow system was simpler to operate and could he operated it a lower cost. 3. That the Callow system operated with a lower air pressure. (The question of the actual power consumption of the two types of machines was not definitely settled at this time.) During the official test period the Callow flotation apparatus showed an additional recovery of from 1 to 1.5 lb. of copper per ton more than the Cole-Bergman machine, this difference being maintained by further treatment of the tailing from both machines. The simplicity of the operation of the Callow flotation system is one of its notable features, two operators per shift being able to handle 90 standard cells and blower equipment.

Jan 7, 1917

By L. Bombardieri, H. F. Mills

Iron King mine, producing gold-silver-lead-zinc-ore, is 10 miles east of Prescott, Ariz. At present the 1806 level is being developed. The echelon pat tern of ore deposit continues at depth but is less pronounced, and the orebody is now more or less continuous over 2600 ft of strike length, with widths 5 ft to as high as 26 ft where overlapping of lenses occurs in one section of the mine. Better continuity of ore at depth has encouraged systematic mining and improved plant facilities to offset higher operating costs. With favorable persistence of ore at depth, the three-compartment No. 6 hoisting shaft appeared inadequate to handle increased tonnage from greater depths, and in 1952, a four-compartment steel and concrete shaft was started. This is now at a depth of 1906 ft, and shaft pockets and loading cartridges are being installed on the 1806 level. Ore from the new shaft, hoisted in 5 ½ -ton Jeto bottom dump skips, is dumped into a 500-ton shaft bin, from which it is fed to a 16x30-in. gyratory crusher and conveyed 415 ft through an undersurface concrete gallery to the secondary crusher and conveying system supplying the 900-ton mill. Lead, zinc, and pyrite concentrates are floated, and the flotation tailings are cyanided after desliming. Waste from the new shaft is dumped into a 250-ton surge bin, fed from this bin to a 24-in. conveyor, and discharged into the glory hole above the block cave. The new shaft is served by a double-drum 400-hp hoist used for ore and waste; the No. 6 shaft conveys men and material.

Jul 1, 1956

By J. Herbst, H. W. Kropholler, L. A. Cramer, E. T. Woodburn, J. C. A. Greene, G. Poling

A generalized mathematical structure is proposed whereby internal species mass flows within a flotation network are defined uniquely by a matrix of enhancement factors. An element of the enhancement factor matrix is defined as the ratio of the mass flow of that species leaving a given cell in the concentrate to the species mass flow leaving the same cell in the tailings. To predict the enhancement factors, knowledge of flotation kinetics effects is required. Recently experimental data has become available relating to the performance of industrial flotation circuits which indicate that the discrete or discrete-distributed kinetic schemes are capable of reasonably effective simulations. These simulations show the importance of including froth effects. The prediction of performance of networks from small-scale batch tests on flotation streams shows that the model for the froth phase used in the simulations does not scale-up well and that aeration effects are also difficult to predict. The network analysis should be useful in differential flotation of complex ores and it is disappointing in this context that little information on the kinetics of activation and depression is available.

Jan 1, 1976