Acid Leaching Process Research Articles

Preparation of High-Purity Nano-Iron Phosphate from Titanium-Extraction Tailings by Co-Leaching Synergetic Ultrasonic-Enhanced Precipitation Process.

The preparation of high-performance electrode materials from metallurgical solid waste is an effective strategy to address current energy and environmental challenges. This study utilizes a mixed acid leaching and ultrasound-assisted precipitation process to extract valuable metallic iron from titanium-extraction tailings (TET) to produce high-purity nano-FePO4 electrode material precursors with unique crystal structures. A leaching efficiency of 95.2% for Fe was attained by using the optimized process parameters, which included a mixed acid concentration of 4 mol/L, a liquid-to-solid ratio of 4:1, and a leaching temperature of 70 °C for 1 h. The optimal precipitation conditions were a pH of 2.0, a temperature of 60 °C, an aging time of 30 min, and a stirring speed of 600 rpm, resulting in FePO4 purity up to 99.6% and fine particle size. Thermodynamic calculations, combined with various characterizations, elucidated the leaching and precipitation mechanisms, highlighting the synergistic effect of phosphoric acid and hydrochloric acid in enhancing the leaching reaction. The thermogravimetric analysis indicated that the decomposition of residual ammonium chloride impurities requires calcination above 360 °C. This research not only provides new insights into the high-value, clean utilization of metallurgical solid waste but also supports sustainable resource recovery and environmental protection by transforming waste into valuable products.

Study on the thermal field material of FZ-Si crystal waste graphite purified by ultrasonic enhanced acid leaching

This study examines the acid-leaching and purification process of waste graphite in the production of Czochralski monocrystalline silicon. The optimal leaching conditions are identified as a liquid-to-solid ratio of 6, a leaching temperature of 70 °C, an acid concentration of 8 mol, and a leaching time of 60 min. The use of hydrofluoric acid, sulfuric acid, and nitric acid in the acid-leaching process increases the fixed carbon content of waste graphite from ∼94 % to ∼98.5 %. To address the low fixed carbon content that cannot be achieved through conventional acid-leaching, a method combining ultrasonic intensification with hydrofluoric and hydrochloric acid leaching is proposed and successfully implemented. Under ultrasonic enhancement conditions, the leaching effect is optimal at a temperature of 60 °C, acidity of 4 mol, and leaching time of 60 min. These results demonstrate that the introduction of ultrasound significantly strengthens the acid-leaching process. The method proposed in this study not only purifies waste graphite through acid-leaching but also elucidates the reaction behavior of various impurity elements during the leaching process. Overall, these findings provide a foundational basis for the recovery of waste graphite in the thermal field.

Enhanced Gallium Extraction Using Silane-Modified Mesoporous Silica Synthesized from Coal Gasification Slag.

This study presents an innovative approach to utilize coal gasification coarse slag (CGCS) for efficient and low-cost gallium extraction. Using a one-step acid leaching process, mesoporous silica with a surface area of 258 m2/g and a pore volume of 0.15 cm3/g was synthesized. The properties of CGCS before and after acid leaching were characterized through SEM, FTIR, XRD, and BET analyses, with optimal conditions identified for maximizing specific surface area and generating saturated silanol groups. The prepared mesoporous silica demonstrated a 99% Ga(III) adsorption efficiency. Adsorption conditions were optimized, and adsorption kinetics, isotherms, and competitive adsorption behaviors were evaluated. Competitive adsorption with vanadium suggests potential application in Ga(III) extraction from vanadium-rich waste solutions. Furthermore, the recyclability of both the acid and adsorbent was explored, with the adsorbent maintaining over 85% adsorption efficiency after five cycles. The adsorption mechanism was further elucidated through SEM-EDS, XPS, and FTIR analyses. This work not only advances resource recovery from industrial waste but also offers a sustainable method for gallium extraction with industrial applications.

Overcoming clay structure challenges in lithium recovery from boron waste using high‐temperature pressure acid leaching

AbstractBoron mines contain significant amounts of lithium along with boron. After boron is extracted, lithium remains in the waste, which has a carbonate‐hosted clay‐type structure, along with other impurities. The scarcity of lithium resources and the increasing need for lithium worldwide make such resources economically important. Although the best hydrometallurgical method for the recovery of lithium trapped within the clay‐structured mineral resources is roasting with chemicals to disrupt the clay structure and acid leaching, the process is quite difficult and costly due to the high energy and chemical addition requirements. To overcome this challenge, this study proposed a high‐temperature–pressure sulphuric acid leaching process to recover lithium from the boron waste. Under the optimized conditions (liquid/solid ratio: 10, acid concentration: 1 M, temperature: 150°C, and contact time: 120 min), 100% of lithium was leached. The leaching mechanism was determined through mineral characterization (X‐ray diffractometry [XRD], X‐ray fluorescence spectrophotometer [XRF], scanning electron microscopy–energy‐dispersive X‐ray spectroscopy [SEM–EDX], Mastersizer), and a shrinking core heterogeneous kinetics model. It was found that high‐temperature–pressure sulphuric acid leaching disrupted clay structure and promoted the leaching of lithium, the leaching kinetics fit the shrinking core heterogeneous kinetics model, and was controlled by a dual mechanism with ash diffusion and chemical reactions on the particle surface. The reaction rate constants increased with increasing temperature, and the activation energy was found to be 32.17 kJ/mol.

Research on the Hierarchical Separation of Rare Earths and Lithium from Rare Earth Molten Salt Slag

Lithium is an important non-renewable resource, and the study of the tiered separation of rare earths and lithium from rare earth molten salt slag is an important approach to solving the current global resource shortage. This article adopts a sulfuric acid leaching process and a lithium-containing solution iron lithium extraction separation process to recover lithium from rare earth molten salt slag. The method systematically investigates the impact of sulfuric acid concentration, liquid-to-solid ratio, leaching temperature, leaching time, pH, P507 concentration, phase ratio, extraction temperature, and extraction time on the lithium extraction effect and iron lithium separation effect in rare earth molten salt slag. The results show that under the conditions of a sulfuric acid concentration of 0.9 mol/L, a liquid-to-solid ratio of 8:1 mL/g, a leaching temperature of 70 °C, a leaching time of 90 min, a pH of 0.9, a P507 concentration of 50%, a phase ratio of 5:3, an extraction temperature of 30 °C, and an extraction time of 20 min, the lithium leaching rate reaches 97.8%, and the separation of iron and lithium is fully achieved. This method can efficiently recover the valuable element lithium from rare earth molten salt slag, which is of great significance for the subsequent preparation of lithium products and the realization of a closed-loop production of rare earth alloys by molten salt electrolysis.

Evironmentally friendly comprehensive utilization of retired waste electrolytes of aluminum electrolysis by calcification leaching

Low energy consumption and environmentally friendly recycling methods can extract high-added value and high-demand elements from retired waste fluorinated aluminum electrolytes (RWE-Al), which is crucial for alleviating resource shortages and environmental hazards. This study aims to develop a method for recovering fluorine and aluminum from RWE-Al through calcium leaching transformation combined with acid leaching. Compared to the traditional high-temperature roasting transformation and high-concentration alkali-metal-salt leaching process, this new process only involves one step of low-temperature leaching, purification of fluorine, carbonization of excess calcium, and evaporation extraction of aluminum, avoiding the generation of high energy consumption and secondary hazardous waste (F-containing liquid/ residue). Fluorine and aluminum are effectively separated and purified with extraction rates exceeding 96%. The effects of leaching temperature, CaO addition coefficient, liquid–solid ratio, and leaching time on the leaching process were systematically studied. The effects of acid concentration, liquid–solid ratio, time, and temperature on the acid-leaching process were also investigated. In summary, the comprehensive utilization of RWE-Al by this new process has proved to be feasible.

Production of hydroxyapatite utilizing calcium and phosphorus in sewage sludge incineration ash

Sewage sludge incineration ash (SSIA), enriching the most of phosphorus (P), could be named as promising P source for P recycling and management. The novel strategy including extraction of P from SSIA by acid (HCl), purification and separation to obtain Ca liquid and Al-P liquid that could produce hydroxyapatite was investigated in this study. The P acid leaching rate was 97.9 % from SSIA at 0.2 mol/L HCl for 2 h. In P purification, P combined with Fe and Al to form precipitation (Al-P and Fe-P) with the increase of pH (<4.0). At a pH of around 4.0, the separation of Ca in the liquid phase was maximized, while Al-P and Fe-P precipitated out. Then, the Al-P precipitation could be dissolved in NaOH solution (0.2 mol/L) to separate Fe-P precipitation, that obtaining the Al-P liquid. What’s more, the high P content (P2O5, 25.5 %) in P product could be produced directly under 1.5 Ca/P molar ratio, which could be used in P fertilizer. The heavy metal content in the P product were far below the Chinese, American and EU fertilizer regulation, indicating that its environment safety. The Ca and P elements leached out during the acid leaching process were eventually recombined to form hydroxyapatite by wet-chemical purification. The cost of P recovery from SSIA was $3.06/kg P product demonstrating commercial potential of this strategy.

Reactivity of a plagioclase concentrate from the South African Bushveld Igneous Complex via extractive acid leaching vs. extractive roasting-leaching processes

This study compared the reactivity of a plagioclase concentrate subjected to two processes: (1) direct acid leaching and (2) thermochemical treatment with ammonium sulfate followed by leaching. The sample was prepared from coarse-grained pyroxenite rock retrieved from the Bushveld Igneous Complex, South Africa. It contained 78% plagioclase (labradorite), 9% orthopyroxene (enstatite) and 13% quartz. The elements contained in the concentrate were categorized into three groups based on their susceptibility to direct acid extraction after 6 h of leaching. Group 1 consisted of the highly reactive main elements of plagioclase (Al, Ca and Na, with extraction efficiencies of 95%, 89% and 81%, respectively). Group 2 included elements predominantly present in enstatite (Mg and Fe with extraction efficiencies of 41% and 55%, respectively). Group 3 was composed of slowly extractable Si (25%) from mostly plagioclase. Increasing the duration of direct acid leaching to 24 h had no effect on the extraction of Group 1 elements, whereas the extraction of Mg and Fe (Group 2) increased to >60%, and that of Si (Group 3) increased from 25 to 80%. The latter correlated with the nearly complete disappearance of the plagioclase blueprint in the XRD pattern of the residues generated after 24 h of leaching. In contrast, plagioclase had limited reactivity with ammonium sulfate during thermochemical treatment. Direct acid leaching of plagioclase-rich tailings can therefore generate leachates to be used as precursors for the preparation of value-added products, such as silica nanoparticles via a sol–gel route and calcium aluminate nanoparticles via solution combustion.

Recovering rare earth elements from svanbergite ores by roasting and phosphoric acid leaching

Rare earth elements (REEs) are regarded as key global strategic metals, and recovering REEs from phosphate ores might constitute an alternative method of acquiring rare earth resources. Svanbergite ores in Sichuan are particular phosphates rocks that contain abundant REEs, which are valuable resources that have not been fully developed and utilized. Consequently, the development of crucial technologies for extracting REEs from svanbergite ores hold strategic and scientific significance. This study employed roasting and phosphoric acid leaching processes to recover REEs from svanbergite ores. Response surface methodology (RSM) was applied to optimize the hydrometallurgical process parameters. This study found that the optimal conditions for REEs recovery were: a roasting temperature 600℃, a roasting time of 38.72 min, a phosphoric acid concentration of 40%, a leaching liquid/solid mass ratio of 10, a leaching temperature of 75℃ and a leaching time of 70 min, resulting in a maximum predicted REE recovery rate of 83.57%. The experimental values exhibited a high degree of proximity to the predicted values, thereby substantiating the accuracy of the model and validating the plausibility of the optimized solution. Mechanistic analysis revealed that after the roasting process, the crandallite formed new acid-soluble new phases. Following the phosphoric acid leaching process, REEs were present predominantly in the form of REE3+ in the leaching solution, consequently separating the REEs from the svanbergite ores. This study provides a valuable reference for the recovery of REEs from associated rare earth phosphate ores.

Sulfuric Acid Leaching of Carbonatite ore with High Calcite Content in Lugiin Gol under Ambient Condition

AbstractCarbonatite bearing rare earth elements are rich sources of rare earth elements and play a critical role in the global supply chain. While carbonatite deposits are less accessible compared to ion adsorption clays, their mineral composition makes them essential for rare earth element extraction. Extraction of rare earth elements from carbonatite ore involves acid leaching, a process that we aim to explore in this study using the carbonatite ore from Lugiin gol. Characterization of the raw material and solid residue post‐leaching was conducted using X‐ray diffractometry, inductively coupled plasma optical emission spectroscopy, and a zeta potential analyzer. These analytical techniques were utilized for kinetic modeling, assessing the influencing factors of leaching, and determining metal recovery in the pregnant solution. The results of the study highlight the significant role of surface charge formation on target particles during milling in the acidic leaching process. This finding shed light on the importance of understanding the surface properties of the particles in the extraction of rare earth elements from carbonatite ores. The surface of particles in the supernatant of rare earth elements’ ore was negatively charged, with the median zeta potential of the 0.025 mm fraction being at its minimum of −60.6 mV.

Recycling of secondary aluminum dross to make alumina by hydrometallurgy: A review

Secondary aluminum (Al) dross (SAD) is a hazardous waste discharged from Al production, processing and recycling. Over 6.7 million tons of SAD was discharged on the planet in 2023. SAD is consisted of 40–60 wt% of alumina (Al2O3), 10–30 wt% of AlN, 5–15 wt% of salts and 3–10 wt% of heavy metal oxides. Currently, recycling of SAD to make Al2O3 by hydrometallurgy is a promising method for disposal of SAD. Hydrometallurgy method is mainly divided into acid leaching and alkali leaching. In acid leaching, Al, AlN and Al2O3 react with acid to form aluminum sulfate and aluminum chloride. In alkali leaching, Al, AlN and Al2O3 react with alkali to form sodium aluminate. High-purity Al2O3 is obtained after precipitation, washing, drying and calcination from the leachate. Resource consumption and emission was calculated to evaluate the economic and environmental benefits. About 147.9 and 172.6 dollars was earned after making Al2O3 from a ton of SAD by alkali and acid leaching process, respectively. And carbon emissions of a ton of Al2O3 was risen about 596.5 and 2216.0 kg CO2, respectively, compared with the Bayer process with bauxite. We proposed a calcination pre-treatment with quicklime on SAD to reduce the carbon emission. The Al and AlN are oxidized into Al2O3 after calcination, and the Al2O3 reacts with CaO to form CaO·Al2O3. The Al in CaO·Al2O3 can be leached out easily with a low concentration of alkali. This review provides a guidance for the recycling of SAD by hydrometallurgy, and proposes a novel idea for the energy and consumption reduction in alumina (Al₂O₃) production.

Integration of CO2 sequestration with the resource recovery of red mud and carbide slag

Red mud and carbide slag, strongly alkaline solid wastes, are produced during the manufacturing of alumina and acetylene gas, respectively. Improper management of these wastes can pose significant environmental hazards. This study proposed a novel method for the collaborative treatment of red mud and carbide slag, enabling the recovery of constituent elements with CO2 sequestration. During the hydrochloric acid leaching process, Si was reclaimed as Si-rich residue, whereas Ti, Al, Ga, Fe, and Sc were stepwise recovered as Ti-rich product, Al/Ga-rich filtrate, and Fe/Sc-rich product via the subsequent hydrolysis, co-precipitation, and alkali leaching processes, respectively. Utilizing carbide slag to replace the energy-intensive NaOH for pH adjustment in the leaching solution resulted in a Ca2+ concentration of 21,922 mg/L in the filtrate after co-precipitation. Through further mineralization process, CO2 sequestration was achieved alongside the production of CaCO3 with the content reaching 98.5 % along with a whiteness of 94.4 %. Furthermore, by adjusting the experimental parameters in the mineralization process, controlled manipulation of the phase and morphology of the CaCO3 product could be achieved to meet various industry requirements. To summarize, the complete resource recovery of red mud and carbide slag was achieved in this study, demonstrating the promising potential in practical applications.

The influence of surfactants on the acid leaching process of a uranium mine in Inner Mongolia

The influence of surfactants on the acid leaching process of a uranium mine in Inner Mongolia

Selective removal of iron from sulfuric acid leaching solution of aerospace magnetic material scraps by jarosite process

Aerospace magnetic material scraps are abundant in cobalt and nickel. Sulfuric acid leaching process is an efficient method for extracting them. But it is a non-selective process, a significant amount of iron dissolves in the solution. This study focuses on the selective removal of iron from this solution using the jarosite process. Eh-pH diagram of K-S-Fe-H2O system was established. Based on thermodynamic analysis, H2O2 is used to oxidize Fe2+ into Fe3+, achieving efficient and selective removal of iron from the solution containing cobalt and nickel. The optimal conditions are as follows: temperature 95°C, K2SO4 dosage coefficient 1.5, seed dosage 10 g/L, time 90 min, pH 1.76, and endpoint pH controlled at approximately 3. Under these conditions, the iron removal efficiency is above 99%, while the loss ratios of cobalt and nickel are below 2%. The product is characterized by XRD and SEM-EDS. Results indicate that the product is jarosite ((K,H3O)Fe3(SO4)2(OH)6), exhibiting an ellipsoid structure with the mean particle size in the range of 0.2–5.0 μm. Temperature, pH value and seed dosage significantly affect reaction rate, particle size and crystallinity, and K2SO4 dosage mainly affects reaction rate and the morphology of jarosite. The jarosite crystallization kinetics can be described by the Avrami equation, with an Avrami index (n) of approximately 2.5 and the apparent activation energy of 42.68 kJ/mol.

Manufacturing of Fe-Ni Composite Oxide Powder and Leaching Process Optimization Via Nickel Pig Iron(NPI) Raw Materials

The lithium-ion battery industry is experiencing rapid growth in response to the widespread adoption of electric vehicles, leading to a significant increase in demand for high-purity nickel compounds, a critical raw material. As a result, ensuring a smooth supply of nickel resources poses a critical challenge not only for the battery industry but also for enhancing the competitiveness of future clean energy industries. High-purity nickel production is predominantly carried out via the high-pressure acid leach (HPAL) process, which extracts nickel from nickel laterite ores. However, this process involves substantial initial capital investment and consumes significant energy due to high temperatures, high pressure, and strong acids. Additionally, this method is associated with environmental concerns such as water and soil pollution. Furthermore, an alternative approach is proposed, which involves adding sulfur to low-grade nickel pig iron (NPI) to produce the nickel matte, followed by an iron removal process to generate high-purity nickel sulfide. However, this method presents challenges such as waste generation from the iron removal process, high carbon emissions, and the production of large amounts of greenhouse gases. This study explored using anodic oxidation to induce oxidative reactions in NPI to nano-powder. This approach improves nickel recovery efficiency by increasing the surface area compared to the bulk state, allowing leaching at room temperature and atmospheric pressure. As a result, the process becomes simpler, with lower initial capital investment, making it economically viable. Additionally, the process requires less energy consumption, allowing for achieving the breakeven point at an earlier stage. This study compared the characteristics of the nano-powderized NPI under different voltage and current ranges. Furthermore, the nickel leaching efficiency was assessed based on leaching solution conditions.

Efficient conversion process and mechanism of bromine in bromine-rich saline wastewater

Bromine resources, widely utilized in basic chemical engineering, are considered non-renewable. A substantial amount of Bromine-Rich Saline Wastewater (BRSW), a hazardous waste, is generated during the smelting process of waste printed circuit boards (WPCBs). The efficient conversion of bromine resources is important in BRSW, which contain high concentrations of bromide ions. This study developed a synergistic treatment method utilizing electrolytic manganese anode sludge (EMAS) achieved efficient conversion of Br− to Br2 in BRSW. Optimal technical parameters are determined through single-factor experiments, and the oxidative acid leaching process was optimized using response surface methodology (RSM). The significant influencing factors for Br− conversion were ranked as acidity > anode mud addition amount > temperature > time, resulting in a bromine conversion efficiency of 96.55%. Kinetic studies of bromine conversion in BRSW during the oxidation-acid leaching process indicated a two-stage process: Stage I being a chemical reaction control model and Stage II being a diffusion control model. The reaction mechanism was elucidated, where H+ initially breaks the Mn-O bond in MnO2, exposing O vacancies and forming Mn4+, leading to a redox reaction between free Br− and Mn4+. This research embodies the concept of synergistic solid waste disposal, achieving cost-effectiveness and streamlined preparation of bromine, and providing technical and theoretical support for bromine resource recovery in BRSW.

Recovery of rare earths from rare earth molten salt electrolytic slag using alkali phase reconstruction method with the aid of external electric field

Rare earth molten salt electrolytic slag (RMES) has emerged as a promising secondary resource for rare earth elements (REEs). This study introduces an innovative leaching technique for extracting REEs from RMES under atmospheric conditions, employing an alkali phase reconstruction method followed by an acid leaching process. Additionally, the external electric field was employed to enhance the reaction. Under the optimal reaction conditions: NaOH initial concentration of 70 wt%, NaOH-slag mass ratio of 4:1, temperature of 160 °C, current density of 1000 A/m2, reaction time of 90 min, stirring speed of 300 r/min, HCl concentration of 4 mol/L, liquid-solid ratio of 15:1, and leaching time of 20 min, the leaching efficiencies of Nd and Pr reach up to 99.21% and 99.14%, respectively. Phase analysis indicates that the rare earth fluorides transform into rare earth hydroxides, significantly enhancing their solubility in acid solution. The imposition of an external electric field leads to pronounced disruption of the RMES surface, thereby promoting the formation of stable reactive oxygen species in the alkaline medium. This facilitates the decomposition of fluorinated rare earths and hastens the phase reconstruction, resulting in an enhanced leaching process. The achieved leaching efficiency with an external electric field is 37% higher than that without an electric field.

Preparation of active zinc oxide from zinc-containing dust and synergistic recovery process of valuable elements

Zinc-containing dust generated from the iron and steel industry is enriched with a large amount of heavy metal substances and rare elements, which is a hazardous waste and a secondary resource. In this study, a new process for the preparation of active zinc oxide and synergistic recovery of valuable elements using zinc-containing dust was proposed. Firstly, a fire process was used to make the valuable components such as Zn, Cd, K, Na, Pb, Bi, In and Sn secondary enriched in the zinc-rich dust. Then the leaching and recovery of K (99.74 %) and Na (98.12 %) were realized by alkaline washing method taking advantage of the solubility difference. Thermodynamic analysis showed that controlling the pH of the leaching system could realize the selective separation of the valuable components. The experimental results showed that the leaching rates of Zn and Cd reached 86.95 % and 92.28 %, respectively, under the optimal experimental conditions, while Pb, Sn, Bi and In were enriched in the neutral leaching slag. Cd sponge (70.3 %) was obtained after the neutral leach solution was treated by Zn powder replacement process. The purified zinc sulfate solution was used to prepare active zinc oxide (97.04 %) by direct precipitation at a molar ratio of CO32–/Zn2+ of 1.15 and a temperature of 50 °C. The average particle size of the product was 21.95 nm, which was in accordance with the industry standard. Neutral leaching slag realized Bi (91.5 %), Sn (71.01 %) and In (94.0 %) in high acid leaching process, and Pb was enriched in slag in the form of PbSO4. The high acid leach solution is hydrolyzed, extracted and replaced to obtain BiOCl, crude tin and crude indium. This process realizes the comprehensive recovery and high-value utilization of various valuable elements in zinc-containing dust sludge, and has significant environmental, economic and social benefits.

An unconventional approach for the efficient recovery of iron, cobalt, copper and silicon from copper slag

High-grade heavy metal elements in copper slag (CS) are worth recovering. Unfortunately, the high viscosity of leaching solution, low leaching efficiency, difficult filtration and low separation efficiency of valuable components exist in the traditional sulfuric acid leaching process. In this study, the above problems are solved by sulfuric acid pretreatment + curing + water leaching. Moreover, iron, cobalt and copper ions in solution are separated by stepwise precipitation. The final iron, cobalt, copper and silicon recoveries are 99.01 %, 98.45 %, 93.13 % and 99.52 %, respectively. Thermodynamic calculations show that H4SiO4 can be converted to insoluble SiO2 to improve filtration properties under curing conditions of sulfur dioxide partial pressures of 10−20∼0 atm, oxygen partial pressures of 10−20∼0 atm and 400–600k. Simulation studies of the phase equilibria of the components of the leach solution by Visual MINTEQ showed that the oxidation of Fe2+ to Fe3+ is necessary for the removal of Fe2+ from the solution by precipitation. This study provides a new idea for the efficient utilization of CS.

Ultrathin carbon film as ultrafast rechargeable cathode for hybrid sodium dual-ion capacitor

The development of electrochemical energy storage devices has a decisive impact on clean renewable energy. Herein, novel ultrafast rechargeable hybrid sodium dual-ion capacitors (HSDICs) were designed by using ultrathin carbon film (UCF) as the cathode material. The UCF is synthesized by a simple low temperature catalytic route followed by an acid leaching process. UCF owns a large adsorption interface and number of additional active sites, which is due to the nitrogen doping. In addition, there exists several short-range order carbons on the surface of UCF, which are beneficial for anionic storage. An ultrafast rechargeable remarkable performance, remarkable anion hybrid storage capability and outstanding structure stability is fully tapped employing UCF as cathode for HSDICs. The electrochemical performance of UCF in a half-cell system at the operating voltage between 1.0 and 4.8 V, achieving an admirable specific discharge capacity of 358.52 mAh·g−1 at 500 mA·g−1, and a high capacity retention ratio of 98.42% after cycling 2500 times at 1000 mA·g−1, respectively. Besides, with the support of ex-situ TEM and EDS mapping, the structural stability principle and anionic hybrid storage mechanism of UCF electrode are investigated in depth. In the full-cell system, HSDICs with the UCF as cathode and hard carbon as anode also presents a super-long cycle stability (80.62% capacity retention ratio after cycling 1300 times at 1000 mA·g−1).